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PUBLISHED    BY 

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METALLURGY 

A   BRIEF   OUTLINE 


OF    THE 


MODERN    PROCESSES   FOR   EXTRACTING 
THE   MORE   IMPORTANT    METALS 


BY 

W.  BORCHERS 

4  t 

KGL.  TECH.  HOCHSCHULE,  AACHEN 


AUTHORIZED  TRANSLATION  FROM  THE   GERMAN 

BY 

WILLIAM  T.  HALL     AND     CARLE  R.  HAYWARD 

Instructor  in  Chemistry  Instructor  in  Metallurgy 

MASSACHUSETTS  INSTITUTE  OF  TECHNOLOGY 


FIRST    EDITION 
FIRST    THOUSAND 


NEW   YORK 

JOHN   WILEY  &    SONS 

LONDON:    CHAPMAN  &  HALL,   LIMITED 
1911 


Copyright,  19x1, 

BY 
WILLIAM  T.   HALL  AND  CARLE    R    HAYWARD 


THE   SCIENTIFIC 
ROBERT  DRUMMOND   AND   COMPANY 
BROOKLYN,    N.   Y. 


PREFACE 


THE  purpose  of  the  book  is  to  present,  in  concise  form, 
the  different  processes  used  for  extracting  the  important 
metals  from  their  ores,  and  for  refining  them.  Both  the 
student  and  the  practical  engineer  can  thus  get  a  broad  view 
of  modern  metallurgical  operations,  without  taking  the  time 
required  for  a  study  of  the  more  detailed  treatises  on  the 
subject.  Whenever  possible,  self-explaining  illustrations  of 
apparatus  have  been  used  in  place  of  descriptions.  The  micro- 
photographs  are  from  articles  by  P.  Goerens,  Aachen,  and 
W.  Campbell,  New  York,  published  in  Metallurgie. 

The  present  translation  is  commended  to  the  public  in  the 
hope  that  such  a  summary  of  metallurgical  processes  will 
prove  of  assistance  to  English-speaking  students  in  pointing 
the  way  to  further  study.  The  translators  are  indebted  especi- 
ally to  Professor  H.  O.  Hofman,  for  his  kind  advice  and  assist- 
ance in  reading  the  proof-sheets. 

WILLIAM  T.  HALL 
CARLE  R.  HAYWARD 


MASSACHUSETTS  INSTITUTE  or  TECHNOLOGY, 
Boston,  Mass.,  February,  1911 


ni 


365552 


TABLE  OF  CONTENTS 


<jOLD 

25 
PLATINUM 

...         28 
SILVER      ..•••* 

56 
MERCURY 

.       •         60 
COPPER     .       •   .    • 

...       •       109 
BISMUTH  . 

.  118 

LEAD          ..•;•' 

.       133 
TIN    .-••••• 

149 

ANTIMONY 

.-160 

NICKEL     .        

177 
IRON          •        •        •        •        •        • 

.        .       218 
CHROMIUM 

.       .       .       •       225 
TUNGSTEN 

.        .        •        229 

CADMIUM 

...       .       .       231 
ZINC 

'»       •       25° 
MANGANESE      .       •       •       ... 

'.       -       253 
ALUMINIUM      ..•••• 


METALLURGY 


GOLD 

Sources 

Natural  Sources: 

NATIVE;  alloyed  with  Pt,  Ag,  Hg,  or  Fe,  as 
VEIN  GOLD,  in  quartz  and  other  ancient  rocks  (primary 

deposits),  and 
PLACER    GOLD,   alluvial    gold   or   gold   dust     (secondary 

deposits). 
MINERALIZED,  rarely;   then  associated  chiefly  with 

TELLURIUM;      also    with     many    sulphides,    particularly 
copper  ores,  whether  as  native  gold  or  as  sulphide,  is 
difficult  to  determine. 
Other  Sources  2 

MATTES  containing  precious  metals,  especially  from  copper 

smelting  and  from  pyritic  smelting. 
ALLOYS  (crude  metals)  from  other  smelting  operations. 
SCRAP  from  the  manufacture  of  jewelry,  and 
OLD  GOLD  and  plated  ware. 

(A)    Extraction 
Ore  Dressing : 

Purely  Mechanical  Dressing,  usually  the  first  method  employed 

for  developing  a  newly-discovered  deposit. 
Electro-magnetic  Concentration,  used  as  a  preliminary  treat- 
ment of  sands  carrying  magnetite. 

Solution  of  Gold  in  Metals:  the  metals  used  as   solvents  are 
copper,  lead,  mercury,  and  silver. 


t    ^« 


*i        »  t          ...»     -  . 


2  METALLURGY 

In  Copper.  The  formation  of  a  gold-copper  alloy  results 
in  most  copper  smelters  when  gold-bearing  copper  ores, 
or  smelter  products,  are  smelted  for  copper;  in  pyritic 
smelting  when  gold-bearing  quartz  is  used  as  lining  for 
blast  furnace,  converter,  or  hearth  of  the  reverberatory 
furnace;  in  the  copper  crust  that  is  removed  when  liquating 
lead  that  carries  precious  metals;  and,  finally,  in  working 
up  old  gold  and  gold-plated  ware,  or  residues  from  the 
manufacture  of  jewelry. 

In  Lead.  A  gold-lead  alloy  is  formed  by  melting  gold- 
bearing  alloys  together  with  lead  ores,  or  metallic  lead, 
and  is  obtained  in  the  subsequent  separation  processes  (sec 
Silver). 

In  Mercury  (Amalgamation).  The  metal  as  it  exists  in  the 
free  state,  or  as  it  is  set  free  from  its  chemical  compounds, 
is  dissolved  in  mercury,  the  resulting  amalgam  is  cleaned 
by  mechanical  treatment,  and  the  precious  metal  is  recov- 
ered by  distilling  off  the  mercury.  Since  silver  compounds 
usually  take  part  in  the  reactions,  an  outline  of  the  process 
will  be  given  under  Silver. 

In   Silver.      A   gold-silver   alloy    is  formed   by    most   of   the 
methods  used   for   treating  ores  containing  silver  and  gold; 
also  by  direct  fusion  preparatory  to  the  parting  of  Au  and  Ag. 
Chemical  Solution  and  Precipitation  of  Gold  : 

Chlorination.  The  free  gold,  as  it  occurs  native  or  as  it  is 
formed  by  a  preliminary  roasting,  is  converted  into  chloride 
by  the  action  of  chlorine  gas  upon  the  moist  ore  (Plattner) 
or  by  the  action  of  dissolved  chlorine  (acid  and  bleaching 
powder)  upon  the  finely  -divided  ore.  The  latter  process, 
as  perfected  by  Thies,  Rothwell,  and  others,  is  the  one  most 
used  in  modern  plants. 

SUITABLE   ORES:     those  which  contain   little    or    none    of 

substances  that  absorb  chlorine  (Cu,  Pb,  Zn,  As,  Sb,  Te, 

S,  CaO,  MgO).     Some  ores  require  a  preliminary  roast. 

The  different  steps  are  as  follows: 

i.  CRUSHING,  accompanied  by  sampling.    The  ore  is  crushed 


GOLD 


to  a  maximum  size  of  60  mm.  and  screened  to  separate 
the  fines  (under  15  mm.)     The  over-size  is  put  through 


FIGS,  i  and  2.  —  Pearce  Turret  Furnace 


rolls  to  reduce  it.     A  sample  is  taken  at  this  point,  and 
the  furnace  charge  is  prepared. 


METALLURGY 


Fin.  5. — American  Chlorination  Barrel  (Scale  i  :  60). 


GOLD 


2.  DRYING  in  a  reverberatory  with  mechanical  stirrers  or 
with  a  rotating  hearth. 


3.  FURTHER  CRUSHING  with  fine  rolls  (7  to  20  meshes  per 
inch). 


METALLURGY 


GOLD  7 

4.  ROASTING  in  Wethey,  Pearce,  or  similar  furnace.    Capacity 
of  the  Pearce  furnace:    100  tons*  of  ore  require  1070  to 
1500  sq.ft.  of  hearth  area,   75  sq.ft.  of  grate    area    and 
10  to   15   tons  of   coal.      Temperature  400°  to    800°    C. 
(750°  to  1550°  F.). 

5.  CHLORINATION  takes  place  in  revolving,  lead-lined  (f  in. 
thick)  barrels  made  of  sheet  iron  (f  in.  thick)  and  provided 
with  a  filter.     Diameter  5  ft.  5  in.  to  5  ft.  7  in.;   length 
14  to  1 6  ft.;    capacity  10  tons  ore  and  5  tons  water.     (Figs. 
3,  4,  and  5.)     For  the  generation  of  chlorine,  200  to  400  Ibs. 
of  H2SO4  and   100  to  200  Ibs.  of  bleaching  powder  are 
required.     Time:   2.5  to  4  hours  with  3  to  5  revolutions  per 
minute.     A  smaller  barrel,  about  4  ft.  7  in.  long  and  3  ft. 
3  in.  in  diameter,  has  a  capacity  of  i  ton  of  ore  and  160 
to  200  gals,  of  water.     For  generating  chlorine  1 8  to  40  Ibs. 
bleaching  powder  and  20  to  50  Ibs.  H2SO4   are  added. 
Twelve  revolutions  per  minute. 

6.  FILTRATION  inside  the  barrel,  and    washing   with    water 
under  2  to  3  atmospheres  pressure  (2.5  to  4  hours). 

7.  PRECIPITATION.     A  preparatory  treatment  with   SO2   to 
react  with  free  chlorine  is  often  used  to-day  when  neces- 
sary.    In    lead-lined    vats    the   gold   is   precipitated   with 
H2S.      (Chlorine    plant,    SO2    and    H2S    generators — see 
Figs.  6-n). 

8.  FILTRATION:    in  filter  presses.    The   filtrate   is   refiltered 
through  sand  filters. 

9.  THE  FILTER  CAKES,  together  with  the  cloths,  are  placed 
in  iron  dishes,  roasted  in  a  muffle  furnace  and  mixed  with 
soda,  borax  and  niter  (Figs.  12-15). 

10.  FUSION  in  graphite  crucibles. 

11.  POURING  into  molds. 

Cyanide  Leaching  and  Precipitation.  This  originated  with  the 
discovery  by  Eisner  in  1844  of  the  solubility  of  gold  in 
alkali  cyanide  solutions,  in  the  presence  of  either  free 

*  The  tons  referred  to  throughout  the  book  are  metric  tons  (2200  Ibs.). 


METALLURGY 


bo 

E 


GOLD  9 

oxygen  or  oxidizing  agents.     According  to  Eisner  the  equa- 
tion is, 

4Au  +  8KCN  +  2H2O  +  O2  =  4KAu  (CN)  2  +  4KOH. 

According  to  Bodlander   (1896)   and  Christy   (1900),  the 
gold-potassium  cyanide    is  formed  by    two    reactions,    viz.  : 


and 

Au2  +4KCN  +  H2O2'=  2KAu(CN)  2  +  2KOH. 

When  the  ore  is  at  all  suitable,  the  cyanide  leaching  proc- 
ess is  the  ideal  supplement  to  amalgamation.  Amalgama- 
tion requires  a  large  grain  that  will  sink  quickly  to  the 
bottom  and  be  taken  up  by  the  mercury  as  the  slime  passes 
through  the  apparatus.  The  cyanide  leaching,  on  the  other 
hand,  requires  finely  divided  gold  for  a  quick  completion 
of  the  solution  process  because,  according  to  the  law  of 
mass-action,  the  speed  of  any  reaction  depends,  up  to  the 
velocity  constant  that  is  characteristic  of  each  reaction,  on 
the  number  of  impacts  of  the  molecules  taking  part  in  the 
reaction,  and  therefore  upon  the  number  of  particles  moving 
in  the  free  state. 

SUITABLE  ORES:  ores  which  are  free  from  substances  that 
destroy  cyanide  or  precipitate  gold  (iron  salts,  organic 
acids,  and  metals  having  a  high  solution  tension  such  as  Cu, 
Zn).  Ores  containing  Te,  As  and  Sb  require  roasting. 
Acid  ores  require  the  addition  of  a  neutralizing  substance 
(milk  of  lime).  Ores  containing  reducing  agents  require 
the  addition  of  oxidizing  agents  (potassium  ferricyanide, 
potassium  permanganate,  peroxides,  bromine,  etc.). 
The  old  method,  which  is  partly  used  to-day,  is  as 
follows,  including  the  preparation  of  the  ore: 

1.  BREAKING  the  ore  in  rock  breakers. 

2.  FURTHER    CRUSHING   and   beginning   of    amalgamation 
in  stamp  mills. 


10 


METALLURGY 


3.  AMALGAMATION  in  shallow  troughs  lined  with  amalgamated 
copper  plates. 

4.  CONCENTRATION    on    bumping     tables.     The    object    of 
this  is  to  separate  out  the  concentrates  (heavy  pyrite  and 

FIG.  16. 


FIG.  18. — Details  of  Iron 
Work. 


FIG.  19. — Details  of  Iron 
Work. 


FIG.  17. — Vat  with  Filter  Bottom  and  Butters'  Distributor. 

large  grains  of  sand).     These,  after  roasting  if  necessary, 

are  treated    with   cyanide   solution  in   the    leaching  vats. 

5.  SEPARATION  OF  THE  TAILINGS  into  sands  and  slimes  by 

means  of  spitzlutten  or  settling  tanks  (Fig.  16-19).     The 


GOLD  11 

spigot  products  from  the  spitzlutten  and  the  settlings  from 
the  settling  tanks  are  treated  in  percolation  vats  and  the 
slimes  in  agitation  vats.  If  the  ore  is  acid,  milk  of  lime 
or  alkali  solution  is  added. 

6.  DRAINING  the  sands  or  concentrates  in  the  percolation 
vats. 

7.  TREATMENT    WITH    STRONG    CYANIDE    SOLUTION    (see 
table  below). 

8.  DRAWING  OFF  of  the  cyanide  solution  and  aerating  for 
several  hours. 

9.  WASHING  WITH  WEAK  SOLUTION:    this  is  repeated  sev- 
eral times  if  necessary,  the  ore  being  allowed  to  aerate  for 
a  short  time  between  treatments. 

10.  WASHING  WITH  WATER,  twice  if  necessary. 

11.  THE  SLIMES  after  the  addition  of  the  cyanide  are  put 
into  vats   (usually  by  means  of  centrifugal  pumps)   and 
agitated  either  with  stirrers,  or  by  blowing  in  air,  in  order 
to  extract  the  gold. 

12.  SEPARATION  OF  THE  SLIME  AND  SOLUTION  by  decanta- 
tion  and  filter-pressing. 

A  summary  of  the  separate  treatment  of  concentrates,  sands 
and  slimes  is  shown  in  the  following  table  j 


Materials. 

Strength  of  Solution. 

Time  of  Leaching. 

Concentrates                     ........ 

o  oi-o  i%  KCN 

2~3    hours 

Sands  

o  01%  KCN 

<—  7    days 

Slimes 

o  01%  KCN 

8—i2  hours 

Recent  Leaching  Methods.  The  higher  extraction  that  is 
possible  from  slimes  and  the  rapidity  with  which  the  gold 
is  dissolved  has  led,  first  in  the  Kalgoorlie  district,  Australia, 
and  subsequently  in  South  Africa  and  in  the  United  States, 
to  sliming  the  tailings,  as  they  come  from  the  stamp  mill  and 
amalgamated  plates,  and  leaching  them  in  this  form.  The 
method  was  carried  out  successfully  in  the  wet  tube  mill  first 


12 


METALLURGY 


I 


GOLD 


13 


built  by  the  F.  Krupp  Co.,  using  the  following  process  as 
developed  by  Diehl: 

1.  PRELIMINARY  CRUSHING  in  rock  breakers. 

2.  FURTHER    CRUSHING    and    beginning    of    amalgamation 
in  the  stamp  mills. 

3.  PLATE  AMALGAMATION. 

4.  FURTHER  PULVERIZING  in  wet  tube  mills.    These  are  rotat- 
ing iron  drums  lined  with  steel  plates  and  filled  with 


Fig.22 


FIG.  23. — Krupp  Wet  Tube  Mill.     Scale,  i :  100. 

rounded  flint  stones.  The  inside  dimensions  of  the 
drums  are:  diameter,  3.5  to  5.5  ft.;  length,  13  to  20 
ft.  Revolutions,  30-24  per  minute.  Power  required, 
24-28  H.P.  Capacity:  One  24-H.P.  mill  will  grind  65 
tons  of  ore  in  24  hours  to  100  mesh  size. 

5.  SEPARATION  OF  THE  SANDS  in  the  spitzlutten  and  carry- 
ing them  back  into  the  tube  mill. 

6.  AGITATION  WITH  CYANIDE,  adding  some  oxidizing  agent 
if   necessary.       By   using  0.2%   KCN  with  0.05%   Br., 


14 


METALLURGY 


GOLD 


15 


Diehl  has  even    leached    telluride  ores  successfully  with- 
out roasting. 

7.  RECOVERY  OF  THE  GOLD  SOLUTION  by  decantation  and 
nitration,  or  by  nitration  alone;    in  filter  presses   (Figs. 


Fig.  28  FILTER  FRAME 


FIG.  29. — Head  Filter  Leaf  with  First  Filter  Frame.     Scale,  i:  15. 

24-34)    Butters'   filters    (Figs.   35-36)    or  Moore's   filters 

(Metallurgie,  4,  559). 

The  precipitation  of  gold  from  the  cyanide  solutions  is 
effected  either  by  clean  zinc  shavings  (old  method),  by 
lead  coated  zinc  shavings  (method  of  Betty  and  Carter),  by 


16  METALLURGY 

electricity  (method  of  Siemens  and  Halske) ,  or  by  electricity 

followed  by  zinc  treatment. 

PRECIPITATION  BY  ZINC  is  accomplished  by  passing  the 
gold  solution  through  long  wooden  boxes,  in  which  the  zinc 
rests  on  filter  bottoms  in  small  cells  with  double  partition 


Fig.  30  FILTER  LEAF 


FIG.  31. — Filter  Leaf  and  Frame.     Scale,  i  :  15. 

walls.  The  solution  rises  through  one  cell,  descends 
between  the  partitions,  rises  through  the  next  cell,  and  so 
on  through  the  vat.  When  the  cells  are  cleaned  up,  the 
larger  pieces  of  zinc  are  removed  mechanically  and  the 
remainder  dissolved  by  treatment  with  sulphuric  acid. 
The  precipitate  is  then  washed,  dried,  melted  with  lead, 
and  cupelled  (see  Cupellation,  under  Silver). 


GOLD 


17 


FIG.  32. 


it.— 1— n-  -'*~fo 4-sUto^"^ =^  r  ~*-- = — -- "-r 


FIG>  23. — Filter  Pressing  Plant  of  the  London  and  Hamburg  Gold  Recovery  Co. 

Ltd.,  London,  Eng. 


18  METALLURGY 

ELECTROLYSIS,  OLD  PROCESS  OF  SIEMENS  AND  HALSKE. 
Anodes:  iron  plates.  Cathodes:  strips  of  lead,  suspended 
in  a  long  wooden  vat  divided  into  cells  by  hollow  partitions. 
The  solution  flows  up  through  the  cells  and  down  through 
the  partitions.  The  vats  are  23  ft.  long,  5  ft.  wide  and 
3  feet  deep;  they  are  divided  into  6  or  8  cells  requiring 
about  100  amperes  total  current  with  a  current  density 
of  about  .05  ampere  per  square  foot  and  a  potential 
of  2  volts  per  cell.  Capacity  1800  cu!ft.  solution  (13,000 
gallons)  per  day.  When  the  deposited  gold  has  reached 


FIG.  35. 

the  proper  thickness,  the  cathodes  are  removed  from  the 
solution,  dried,  melted,  and  cupelled  for  crude  gold  (850-900 
fine). 

ELECTROLYSIS  BY  MORE  RECENT  METHODS.  For  solutions 
low  in  copper,  Butters'  process  may  be  used.  The  cathodes 
are  of  tin  plate,  from  which  the  gold  continually  drops 
off  to  the  bottom  of  the  precipitation  vat,  which  is  built 
like  a  spitzlutte,  and  thence  it  is  withdrawn  by  means 
of  a  small  stream  of  liquid  and  filtered  outside  the  vat. 
Since  with  solutions  running  high  in  copper  the  deposit 
is  more  dense,  Charles  P.  Richmond  (Metallurgie  4,  502) 


GOLD 


19 


has  gone  back  to  lead  cathodes.     These,  when  covered 
with  a  gold  or  copper  deposit  of  proper  thickness,  are 


FIG.  36. — Butters'  Vacuum  Filter.  Frames  made  of  iron  rods  covered  with 
canvas.  The  spaces  between  are  filled  with  cocoa  matting.  The  perforated 
iron  pipes  are  connected  to  suction  tubes. 

Operation  of  the  filter:  The  vat  is  filled  with  slime  and  suction  applied. 
The  solution  is  drawn  through  the  hollow  frame  until  the  cake  on  the  outside  of 
the  filter  becomes  so  thick  that  filtration  is  difficult.  The  slime  is  then  withdrawn 
from  the  vat  and  water  admitted  for  washing  the  cakes.  After  washing,  the  cakes 
are  removed  from  the  filter  by  applying  pressure  instead  of  suction. 


FIG.  37. 
Cross-sections  of  Zinc  Boxes. 


FIG.  38. 


removed  from  the  solution,  enclosed  in  filter  bags,  and 
placed  in  an  acid  bath,  where  they  are  used  as  anodes. 
The  gold  remains  behind  as  slime,  while  copper  is  dissolved 


20 


METALLURGY 


and  goes  to  the  cathode,  but  if  a  high  current  density  is 
used  it  falls  off  and  is  recovered  as  a  slime  similar  to  cement 


FIG.  39.— Plan. 


A-B 


FIG.  40. — Sections  EF  and  CD. 


J°]J 


FIG.  41. — Section  AB. 


copper.     The  gold  which  remains  as  slime  in  the  bags 
is  dried  and  fused  as  described  above. 
Parting :  gold  as  a  residue.  Most  of  the  so-called  parting  processes 


GOLD 


21 


consist  in  dissolving  away,  either  in  aqueous  solutions  or  by 
fusion,   the  metals  alloyed  with  the  gold,   leaving  the  latter 


FIG.  42. — Anode. 


FIG.   43. — Cathode:    in  the      FIG.  44. — Wooden 
Second  Part  of  the  Process:  Frame. 

Anode. 


behind.  Of  the  many  methods  that  have  been  devised,  the 
following  are  still  used  in  practice : 

Nitric  acid  parting,  inquartation         j 

Sulphuric  acid  parting,  \  Described  under  Silver. 

Fusion  with  sulphur.   Rossler  process  J 

Electrolytic  parting  and  purification  of  other  metals.  See 
Silver,  Copper  and  Nickel. 


(B)  Gold  Refining 

Solution  and  Precipitation  of  the  Gold : 

Chlorinating  Solution.  Chlorine  and  solutions  evolving  chlorine, 
particularly  aqua  regia,  are  the  principal  agents  used 
in  the  final  refining  of  gold.  For  dissolving  the  gold,  dilute 
aqua  regia  suffices  (2  parts  concentrated  HC1,  i  part  concen- 
trated HNOs,  3  parts  H2O).  Apparatus:  porcelain  vats. 
The  gold  solution  is  filtered  to  remove  the  insoluble  residue 
(AgCl)  and  the  following  operations  are  then  carried  out: 
Precipitation  of  the  gold  by  ferrous  sulphate,  filtration, 
washing,  drying,  and  melting  in  a  graphite  crucible  under 
a  cover  of  glass  and  borax.  If  platinum  is  present,  it  is 


22  METALLURGY 

precipitated  by  adding  metallic  iron  to  the  solution  from 

which  the  gold  has  been  removed. 
Electrolytic  Parting,  used  for  refining  gold  that  contains  platinum. 

ANODES:   gold  platinum  alloy. 

CATHODES:  sheets  of  refined  gold.  Distance  between  elec- 
trodes 1.2  inches. 

ELECTROLYTE:  gold  chloride  solution  containing  25-30  oz. 
of  gold  and  20-50  oz.  HC1  (sp.gr.  1.19)  per  cubic  foot 
(7.5  gallons).  If  the  anodes  contain  lead  some  H2SO4 
is  added.  The  large  amount  of  free  HC1  in  the  electro- 
lyte is  required  because  the  anode  gold  goes  into  solution 
only  when  the  conditions  are  favorable  for  the  formation  of 
HAuCU,  the  reaction  being:  Au  +  3Cl  +  HCl  =  H(AuCl4). 
Furthermore,  the  greater  the  current  density,  the  larger 
the  quantity  of  free  acid  that  must  be  kept  in  the  bath; 
a  high  current  density  is  required  for  the  rapid  de- 
position of  the  gold.  Electrolytic  solution  of  the  gold  is 
increased  by  rise  of  temperature.  The  conditions  for 
electrolysis  are,  therefore: 

CURRENT  DENSITY:  100  amperes  per  square  foot,  at  times 
rising  as  high  as  300  amperes  per  square  foot. 

POTENTIAL:    i  volt. 

TEMPERATURE:  60°  to  yo°C.  (140°  to  158°  F.). 

ELECTRO LYZING  VESSELS:  stone  jars  or  porcelain  beakers 
immersed  in  a  water  bath.  An  automatic  method  is  used 
for  replacing  the  water  lost  by  evaporation  from  the  water 
bath;  the  wash- water  from  the  cathodes  and  anode  mud 
is  used  chiefly  for  this  purpose. 

PLATINUM,  although  insoluble  in  HC1  when  used  alone  as  an 
anode,  goes  into  solution  somewhat  in  the  presence  of  gold, 
but  it  is  not  deposited  on  the  cathode  as  long  as  the  amount 
of  platinum  in  the  electrolyte  does  not  become  more  than 
double  the  amount  of  gold.  The  platinum  may  be  precipi- 
tated from  a  platinum-rich  electrolyte  by  adding  NH4C1, 
and  the  same  is  true  of  palladium.  Silver  is  converted 
into  insoluble  AgCl. 


GOLD  23 

THE   PRODUCTS   of   electrolytic   refining   are  gold,  999.8  to 

1000  fine,  platinum,  and  silver  chloride. 
Dissolving  or  Slagging  the  Metals  Alloyed  with  the  Gold: 

SULPHURIC  ACID  REFINING  consists  in  repeatedly  boiling 
the  gold  with  concentrated  H2SO4.  The  further  treat- 
ment of  the  gold  residue  is  the  same  as  described  under 
Chlorinating  Solution  (p.  21). 

NITRIC  ACID  REFINING  consists  in  boiling  repeatedly  with 
concentrated  nitric  acid  (see  Silver,  Ag-Cu  parting). 


FIG.  45. — Crystals  of  Precipitated  Gold  (X35). 

REFINING  FUSION  :  gold  containing  small  amounts  of  impur- 
ities, and  also  the  precipitated  gold  as  described  on  p.  21, 
which  is  to  be  melted  and  cast  into  bars,  is  fluxed  with 
borax  and  glass,  and  if  necessary  with  an  oxidizing  agent 
such  as  niter,  and  melted  in  graphite  crucibles. 
Properties  of  Refined  Gold  : 

SPECIFIC  GRAVITY:    19.3. 

COLOR:   yellow  with  brilliant  lustre. 

TENACITY:  very  tough,  the  most  tenacious  of  metals. 

FRACTURE:   hackly. 


24  METALLURGY 

STRUCTURE:   (Fig.  45). 

MELTING  POINT:   1064°  C.  (1947°  F.). 

VAPORIZATION  at  2000°  C.  (3632°  F.). 

ELECTRICAL  CONDUCTIVITY:  0.6  to  0.7  referred  to  Ag=i. 

ALLOYS  with  most  metals.    The  alloys  of  Au  with  Pt,  Ag,  Hg, 

Cu,  Pb,  and  Zn  are  important  metallurgically. 
CHEMICAL   BEHAVIOR:    not   very   active,  has   a  low  solution 

tension,  and   its   compounds  are  readily  broken  down.     In 

extracting  gold,  the  most  important  solvents  are  Cl,  Br,  and 

KCN. 


PLATINUM 

Sources 

Natural  Sources : 

FREE,  alloyed  with  the  other  platinum  metals  as  well  as  with 

Fe,  Au,  and  Cu. 

MINERALIZED,    occasionally   as   a   compound   with   arsenic 
(Sperrylite,  PtAs2)  but  more  frequently  in  Ni,  Cu,  and  Au 
ores,  although  in  small  quantities. 
Other  Sources : 

Smelter  products:    In    matte    and    bullion    from    smelting 
platinum-bearing  Ni,  Cu,  and  Au  ores. 

Extraction 

Concentration.    Wet,  mechanical  methods  are  nearly  always  util- 
ized in  the  preliminary  concentration  of  platinum-bearing  gravel. 
Solution  in  Metals : 

In  Lead.    This  is  carried  out  occasionally  by  smelting  platinum 
ores  with  PbO  and  PbS,  whereby  the  Fe  and  Cu  in  the  ores 
are  converted  into  matte  while  the  platinum  metals  enter  the 
lead  bullion,  from  which  they  are  subsequently  recovered  by 
cupellation  (see  Silver  Cupellation  Methods,  pp.  33-35). 
In  Nickel   ]  These  processes  take  place  along  with  the  treat- 
In  Copper  V     ment  of  Ni,  Cu,  and  Au  ores.      See  these  metals 
In  Gold      J      and  their  electrolytic  separation. 
Chemical  Solution  and  Precipitation  of  the  Platinum : 

Chlorination  in  aqueous  solution  by  means  of  mixtures  evolving 
chlorine,  especially  aqua  regia,  is  carried  out  in  the  refining 
of  alloys.  From  the  resulting  solution  the  platinum  is  sub- 
sequently precipitated  either  by  iron  (see  Pt-Au  parting), 
or  by  ammonium  chloride. 

25 


26  METALLURGY 

Refining 

Chemical  Solution   and  Precipitation  is  utilized  both  in  the 

direct  working  of  the  ores  and  in  the  parting  of  alloys  or  of 

crude  platinum.     The  following  treatments  are  given: 

Purification  with  HC1,  containing  but  a  small  amount  of  HNO3, 

so  that  only  the  base  metals  are  dissolved.     This  treatment  is 


FIG.  46. — Surface  of  Fused  Platinum  (X33). 

not  given  unless  imperative.  As  a  rule  it  is  only  necessary  to 
carry  out 

Chlorination  in  aqueous  solution  as  described  above.  If  the 
solution  carries  gold,  this  can  be  precipitated  by  ferrous 
sulphate  or  by  electrolysis.  The  separation  of  the  accom- 
panying platinum  metals  takes  place  usually  by 

Partial  Precipitation  with  NH4C1,  if  necessary  after  the  pre- 
vious reduction  of  Pd  and  Ir  chlorides  from  the  "  ic  "  to 
the  "  ous "  state.  Ammonium  chloride  precipitates  the 
platinum  as  ammonium  chloroplatinate,  (NH4)2PtCl6.  The 
latter  after 


PLATINUM  27 

Separation  from  the  Solution    by  decantation,  filtering,  wash- 
ing, and  drying  is  changed  by 
Ignition  into  platinum  sponge: 

(NH4)  2PtCl6  =  Pt  +  2NH4C1  +  2C12. 

Melting.     In  order  to  convert  the  precipitated  powder  or  the 
sponge  into  a  compact  mass,  it  is  heated  by  an  oxyhydrogen 
blowpipe     in    a     furnace    made   of   limestone.     This    also 
removes  the  last  traces  of  impurities. 
Properties  of  Platinum : 
SPECIFIC  GRAVITY:  21.5. 
COLOR:  grayish  white,  brilliant  lustre. 
TENACITY:    very  ductile,  very  high  tensile  strength,  capable 

of  being  drawn  into  the  finest  wires  and  the  thinnest  sheets 

(platinum  foil). 
STRUCTURE  :  see  Fig.  46. 
MELTING-POINT:  1745°  C.  (3173°  F.). 
ELECTRICAL  CONDUCTIVITY:  0.08  (Ag=i). 
ALLOYS   readily   with   most  metals,  especially   when   easily 

fusible,  also  with  H. 
CHEMICAL  BEHAVIOR:  not  very  active,  low  solution  tension; 

compounds    easily    dissociated.      For   the    extraction   of 

platinum  Cl  is  the  most  important  solvent. 


SILVER 

Sources 

Natural  Sources : 

NATIVE,  alloyed  with  Au,  Cu,  Hg. 

MINERALIZED,  principally  in  combination  with  the  halogens 
and  sulphur.  Hornsilver,  AgCl.  Bromargyrite,  AgBr. 
lodargyrite,  Agl.  Silver  glance,  Ag2S,  occurs  free,  in 
sulpho  salts  (ruby  silver,  tetrahedrite),  and  as  solid  solu- 
tion in  sulphides. 

Other  Sources : 

Burnt  pyrite,  matte  from  copper  and  lead  smelting,  slags, 
drosses,  and  alloys  from  metallurgical  processes  (black 
copper,  base  bullion,  zinc  skimmings,  etc.). 

OLD  METAL  and  metallic  scrap. 

(A)    Extraction 

Concentration:  Nearly  always  carried  on  in   connection    with 

amalgamation  (p.  35  et  seq.). 
Solution  of  Silver  in  Metals :  Copper,  lead,  and  mercury  are 

commonly  used  as  solvents. 
IN  COPPER:  (see  Gold  and  Copper). 

IN  LEAD:  Silver,  as  it  occurs  free,  or  as  it  results  from  the 
decomposition  of  its  compounds,  is  taken  up  by  molten  lead, 
and  the  two  metals  are  subsequently  separated.  Materials 
poor  in  silver  (under  10%)  are  smelted  with  lead  ores  (see 
Lead).  Materials  rich  in  silver  (over  10%)  are  treated  on 
a  bath  of  molten  lead  (see  Cupellation,  pp.  33-35). 

With  lead  as  the  carrier,  most  sulphide  ores  are  satisfac- 
torily treated,  either  directly  or  after  roasting.  Lead-free 
copper  ores  carrying  precious  metals  are  not  used  for  this 

28 


SILVER 


29 


purpose,  but  are  worked  by  treating  with  them  copper  acting 
as  the  carrier.     Ores,  either  raw  or  roasted,  containing  anti- 
mony and  arsenic  and  various  other  materials  not  too  rich  in 
copper  are  used. 
Operations : 
i.  SMELTING  of  materials  low  in  silver,  with  lead  ore  to 


FIG.  47. — Refining  Furnace.    Scale,  i :  100. 

obtain  a  base  bullion  low  in  silver  (about  i%  Ag).      Smelt- 
ing directly  to  rich  bullion  leads  to  high  slag  losses. . 
REFINING  THE  BASE  BULLION.    Melting  in  a  deep-hearth 
reverberatory  furnace  at  a  low  temperature.     Removal  of 
the  mechanically-enclosed  impurities  and  of  the  difficultly- 
fusible  constituents ;  the 
material       withdrawn, 
usually  copper-bearing, 
is  returned  to  the  ore- 
smelting  furnace. 

Heating  in  an  oxidiz- 
ing flame  at  a  high  tem- 
perature in  order  to  aoo 
remove  the  larger  part  of 
the  antimony.  The  oxi- 
dation product  (skim- 
mings: Pbs(SbO4)2  + 
#PbO),  is  refined  under 
a  covering  of  charcoal  in  order  to  reduce  part  of  the  PbO. 
The  refined  skimmings  are  smelted  for  hard  or  antimonial 
lead,  a  Pb-Sb  alloy  with  14-20%  Sb. 


0 

, 

\ 

E 

A 

B     \C 
\ 

303°C 

R 

ft 

-4*-^ 

D                   5                       10 

FIG.  48. 

30  METALLURGY 

3.  CONCENTRATION  of  the  silver  in  a  portion  of  the  lead. 
This  may  be  accomplished  either  by  crystallization  (Pattin- 
son  process)  or  by  crystallization  after  adding  zinc,  and  then 
separating  the  Ag  as  an  Ag-Zn-Pb  alloy  (Parkes  process). 
The  enrichment  of  the  bullion  during  the  crystallization 
of  the  lead  takes  place  according  to  the  accompanying 
freezing-point  curve  of  silver-lead  alloys.  Since  the  eutectic 
contains  about  4%  silver,  on  cooling  a  bullion  with  about 
i%  silver,  from  340°  C.,  lead  will  begin  to  separate 
at  E,  thus  enriching  the  mother-metal  in  silver.  If  we  cool, 
for  example,  to  306°  C.,  the  ratio  of  crystallized  lead  to  the 
still  liquid  silver-lead  alloy  will  be  as  BC:AB.  In  prac- 
tice it  is  not  possible  to  accomplish  such  a  sharp  separa- 
tion between  Pb  and  Pb-Ag  alloy,  for,  in  the  skimming  off 
of  the  crystallized  lead  or  in  the  withdrawal  of  the  liquid 
Pb-Ag  alloy,  some  of  the  alloy  adheres  mechanically  to 
the  crystals,  or  small  crystals  remain  suspended  in  the 
molten  alloy.  Two  methods  have  been  devised  for  the 
practical  working  of  this  process: 

THE  ORIGINAL  PATTINSON  PROCESS.  The  base  bullion 
is  melted  in  iron  kettles  and  the  dross  skimmed.  The 
cooling  of  the  lead  is  hastened  by  spraying  with  water. 
After  crystallization  begins,  the  crystals  are  transferred 
by  means  of  a  perforated  iron  skimmer  to  a  neighboring 
kettle  and  the  process  continued  until  only  about  one- 
third  of  the  original  contents  remains  in  the  first  kettle. 
To  the  remaining  liquid,  either  in  the  same  or  a  neighbor- 
ing kettle,  more  lead  of  the  same  silver  value  is  added 
and  the  process  repeated.  Similarly  the  skimmed 
crystals,  which  are  not  yet  sufficiently  desilverized,  are  re- 
melted  and  subjected  to  the  same  treatment.  Thus  the 
kettle  at  one  end  of  a  row  will  finally  hold  the  enriched 
bullion  (1.5  to  2%  Ag)  while  the  kettle  at  the  other  end 
will  receive  the  desilverized  lead. 

THE  LUCE-ROZAN  PROCESS  (Figs.  49-50,  p.  32).  The 
base  bullion  is  melted  in  tilting  kettles  which  discharge 


SILVER  31 

into  the  crystallizing  kettle.  Steam  is  blown  through 
the  latter  until  two-thirds  of  the  lead  has  been  crystallized 
and  then  the  enriched  bullion  is  removed  by  tapping.  The 
crystallizing  kettle  is  heated  to  remelt  the  lead  crystals  and 
refilled  with  bullion  of  the  same  silver  value.  The  crys- 
tallization is  repeated  and  the  process  continued  until 
the  lead  remaining  in  the  kettle  is  sufficiently  desilverized. 
It  is  remelted  and  tapped  into  molds. 

ZINC  DESILVERIZATION  is  based  upon  the  slight  solubility 
of  Zn  in  Pb,  and  conversely  of  Pb  in  Zn,  and  upon  the 
high  melting-point  of  the  Zn-Ag-Pb  alloy  compared  with 
that  of  the  lead. 

At  350°  C.,  Pb  dissolves  0.6  Zn  and  at  650°  it  dissolves 
3%  Zn. 

The  different  operations  of  zinc  desilverization  are  as 
follows : 

The  Cu  is  removed  from  the  previously-purified,  melted 
base  bullion  by  adding  to  it  a  small  amount  of  Zn  (2%), 
allowing  it  to  cool  and  removing  the  crystallized  alloy. 
As  soon  as  a  sufficient  quantity  of  copper  skimmings  have 
accumulated,  they  are  liquated,  thus  producing  (i)  liquated 
bullion,  which  is  returned  to  the  bullion  that  has  already 
been  freed  from  Cu,  and  (2),  a  copper  alloy  low  in  silver 
(copper  dust) .  The  copper  dust  is  melted  and  oxidized 
with  steam,  thus  separating  it  into  a  rich  bullion  and 
oxides  of  Cu  and  other  metals.  These  rich  oxides  are 
leached  with  H2SO4  so  that  a  solution  containing  the 
sulphates  of  copper  and  of  zinc  results  and  a  residue  com- 
posed of  particles  of  bullion  which  had  been  entangled  in 
the  oxides.  The  copper  is  precipitated  from  the  solution 
by  zinc  scrap  and  the  resulting  solution  of  zinc  sulphate 
yields  zinc  vitriol,  ZnSC^.yH^O,  on  evaporation. 

The  bullion,  freed  from  copper,  is  now  heated  up  and 
the  greater  part  of  the  silver  removed  by  means  of  a 
larger  addition  of  Zn  (in  the  form  of  pure  zinc,  and  skim- 
mings poor  in  silver).  The  alloy  which  has  crystallized 


32 


METALLURGY 


FIG.  50. — Pattinson  Plant,  using  the  Luce-Rozan  Process. 


SILVER  33 

out  (zinc  crust),  is  removed  and  liquated  in  another  kettle. 
The  liquated  lead  goes  back  to  the  desilverizing  kettle. 
The  residue  (rich  zinc  crust)  is  distilled,  giving  rich  bullion 
and  zinc. 

The  desilverized  lead  is  freed  from  zinc  by  blowing 
steam  through  it.  The  resulting  mixture  of  zinc  oxide 
and  lead  oxide  which  floats  on  the  surface  of  the  metal 
is  skimmed,  washed,  and  sold  as  pigment.  If  the  lead 
still  contains  antimony,  further  oxidation  is  accomplished 
by  means  of  an  air  current.  The  antimony  is  thus  removed 
in  the  litharge  which  is  formed  at  the  same  time.  As  a 
residue,  there  remains  in  the  kettle  pure  lead,  the  so-called 
"soft  lead." 

As  apparatus  for  desilverizing,  hemispherical  or  flat 
cast-iron  kettles  with  stirrers  are  used;  for  the  distillation 
of  the  rich  zinc  crust,  graphite  or  clay  retorts;  and  for 
cupelling,  reverberatories  with  hearths  made  of  bone-ash 
or  cement  and  lime.  .  Lead-lined  wooden  vats  serve  for 
working  up  the  so-called  rich  oxides,  and  for  the  poor 
oxides  resulting  from  the  dezincification  of  the  lead,  a  jig 
is  used  together  with  a  series  of  wooden  slime  tanks. 

The  rich  bullion  is  converted,  by  cupelling,  into  litharge 
and  crude  silver,  both  the  English  and  German  methods 
being  used.  In  the  German  process  large  reverberatory 
furnaces  are  used  which  take  the  entire  charge  at  the 
beginning  of  the  operation.  In  the  English  process,  small 
reverberatory  furnaces  are  used,  the  hearths  being  kept 
full  during  the  operation  by  repeated  additions  of  bullion 
until,  finally,  they  are  full  of  crude  silver. 

If  liquated  bullion  is  not  used,  the  first  litharge  drawn 
off  will  be  impure.  After  the  removal  of  this,  silver  scrap 
may  be  added  to  the  bullion.  The  base  metals  are 
oxidized  and  the  precious  metals  dissolved  in  the  bullion. 
Now  begins  the  cupellation  proper,  namely  the  oxidizing 
smelting  of  the  lead  whereby  the  resulting  litharge  is  con- 
stant! v  allowed  to  flow  out  of  the  furnace.  Part  of  the 


34  METALLURGY 

litharge  is  absorbed  by  the  hearth  which,  before  starting, 
has  been  lined  with  bone-ash  (now  little  used)  or  cement 
and  lime.  During  the  cupellation  the  metal  surface  is  com- 
pletely covered  with  yellow,  glowing  litharge  which  at  the 
last  recedes  from  the  bluish  surface  of  the  metallic  silver 


FIG.  51. — Desilverizing  Kettle  and  Auxiliary  Apparatus,  with  stirrer, 

(the  blick).  Since,  in  spite  of  the  fact  that  the  surface  of 
the  lead  is  covered  with  litharge  during  almost  the  whole 
of  the  cupelling  operation,  the  oxidation  takes  place  com- 


FIG.  52 — Desilverizing  Kettle  and  Auxiliary  Apparatus,  with  siphon. 

paratively  quickly,  it  must  be  ascribed  to  the  formation 
of  a  higher  oxide  of  lead,  PbC>2.  In  fact,  PbC>2  is 
formed  readily  in  an  oxidizing  atmosphere  at  the  temper- 
ature of  the  cupelling  hearth  in  the  presence  of  strongly 
basic  oxides.  PbO,  however,  is  itself  a  strongly  basic 


SILVER 


35 


Desilverizing  Kettle  and  Auxiliary  Apparatus. 

Fig.  51,  with  stirrer.    Fig.  52,  with  siphon.    Figs.  53  and  54,  with  lead  pump  and 

molds. 

oxide,  and  can  form  compounds  with  PbO2  such  as,  for 
example,  the  lead  plumbate  found  in  red  lead : 


36  METALLURGY 

This  compound  carries  the  oxygen  of  the  air  through 
the  litharge  layer  to  the  lead.  Without  it,  the  oxidation 
of  the  lead  would  take  place  much  more  slowly,  or  some 
device  other  than  the  strong  air  current  playing  upon 
the  melt  would  be  necessary  to  keep  fresh  surfaces  of  lead 
in  contact  with  ah*. 

The  products  from  the  true  cupellation  are:  litharge, 
PbO;  hearth  lining  saturated  with  litharge;  lead  fume, 
the  volatilized  lead  oxide  which  has  been  caught  in  the 
dust  chamber  and  contains  other  volatile  substances; 
and  finally  the  silver  itself,  which  is  known  as  blick  silver. 
This  last  may  contain  small  amounts  of  lead,  bismuth, 
copper,  etc.,  also  any  other  precious  metals  that  may  have 
been  present  in  the  ores.  The  litharge,  hearth,  and  flue 
dust  go  back  to  the  ore  smelter. 

Amalgamation.  This  includes  the  solution  in  mercury  of  pre- 
cious metals,  either  native  or  resulting  from  the  decomposition 
of  their  compounds,  the  mechanical  purification  of  the  result- 
ing amalgam  and  the  separation  of  the  precious  metals  from 
the  mercury  by  distillation. 

AMALGAMATION  WITHOUT  CHEMICALS:  applicable  to  ores 
containing  the  precious  metals  in  the  free  state  or  as  com- 
pounds decomposable  by  mercury.  It  is  assumed  in  the 
first  case  that  the  precious  metals  are  not  widely  dissem- 
inated through  the  gangue.  Coatings  of  other  metallic 
compounds  surrounding  the  particles  of  precious  metals 
may  also  prevent  the  necessary  contact  of  the  latter  with 
the  mercury.  The  following  methods  are  used  in  this 
type  of  amalgamation: 

AMALGAMATION  DURING  CRUSHING  AND  SUBSEQUENT 
MECHANICAL  TREATMENT:  These  methods  differ  from 
ordinary  ore-dressing  processes  only  by  the  use  of 
mercury,  free  or  in  the  form  of  copper  or  silver  amal- 
gam, in  various  parts  of  the  apparatus  used  for  crush- 
ing and  concentrating;  in  the  last  case,  silver-plated 
copper  plates  are  used.  The  apparatus  of  the  ore- 


SILVER 


37 


SECTION  a-&,  c-d 


"•'•-• 


Fig,  55 


FIG.  56. — German  Cupelling  Fumace. 


38  METALLURGY 

dressing   plants   requires    no  important  changes,  so  that 
the  following  divisions  may  be  made  of  these    amalga- 
mation processes: 
HYDRAULIC  METHODS  in  connection  with  sluice  amal- 


FIG.  59. — English  Cupelling  Furnace. 

gamation   (Hg  is  kept  on  the  bottom  of  the  sluice 
in  grooves  of  the  pavement). 

STAMP-MILL  AMALGAMATION       1      Both    processes    are 

WET-GRINDING  AMALGAMATION  J  completed  on 

AMALGAM  CATCHERS  AND  AUXILIARY-AMALGAMATORS. 

Amalgamated  copper  plates  used  as   the  bottom  of 

shallow  sluices    or    frequently,    amalgamating    pans; 

as  for  example  in  the  Laszlo  system,  the  tailings  from 

which  it  is  sometimes  necessary    to  concentrate   by 

passing  over  shaking-tables  of  the  Frue,  Wilfley,  or 

Ferraris  type,  especially  for  recovering   pyrite,   arsen- 

opyrite,  etc. 


SILVER 


39 


CRUSHING   AND    AMALGAMATION  AS  SEPARATE  PROCESSES. 
In  this  case  the  amalgamators  serve   chiefly  for  mixing 


FIG.  60. 


FIG.  61. — Laszlo  Amalgamator. 

the  ore  with  mercury.    They  may  be  classified,  according 
to  their  form  into 
MORTAR  AMALGAMATION. 


40  METALLURGY 

WET-MILL  AMALGAMATION  OR  PAN  AMALGAMATION: 
Schemnitz  amalagmator,  Laszlo  amalgamator,  Ameri- 
can amalgamating  pans.  These  contain  amalgamated 
copper  plates  as  bottoms  to  shallow  sluices,  and 
various  sluices  with  devices  for  feeding  the  mercury 
and  collecting  the  amalgam. 

AMALGAMATION   WITH   CHEMICALS:    the  ores  may  be  amal- 
gamated without  previous  roasting  by  the 

Patio  Process  or  Old-fashioned,  Mexican  Heap-Amal- 
gamation. This  consists  of  the  following  operations: 


FIG.  62. 


Arrastra. 


FIG.  63. 


1.  PRELIMINARY  CRUSHING  of  the  ore  in  edge  mills 
or  stamp  mills. 

2.  FURTHER  CRUSHING  in  wet  grinders  called  arrastras 
(Figs.  62  and  63). 

3.  PARTIAL   DE- WATERING   OF   THE    SLIMES   in   low 
heaps  surrounded  by  sand  dams. 

4.  PREPARATION  OF  THE  AMALGAMATION  HEAPS  (tor- 
tas)    by    spreading    out    the    now-thickened    slime 
upon  a  paved  floor  surrounded  by  stones.    Diameter 
of  the  floor:    25  to  50  ft.,  depth  6  to  12  in. 

5.  INCORPORATION    (by   treading  in   with   mules)    of 
from  2  to  5%  of  common  salt  (based  on  the  weight 
of  ore);   0.25  to  0.5%  of  CuSO4.5H2O,  in  the  form 
of  magistral   (a  product  obtained  by  the  roasting 


SILVER  41 

of  pyrite  which  contains  CuSO4  and  CuCl2  besides 
ferric  salts);  and  6  to  8  Ibs.  of  Hg  per  pound 
of  Ag  in  the  ore. 

6.  WET  CONCENTRATION  OF  THE   SLIMES  after  the 
completion  of  the  amalgamation. 

7.  TREATMENT  OF  THE  AMALGAM  (see  below). 

The  Patio  process,  in  spite  of  its  great  age  (already 
in  use  300  years)  is  a  very  slightly  developed   method 
which  gives  satisfactory  results  only  with  ores  that  have 
been  prepared  by  weathering  processes,  but  the  mercury 
loss  is  always  high  (200%  of  the  recovered  silver).    For 
the  chemical  reactions  see  the  Krohnke  process. 
KROHNKE  PROCESS:     suitable  for   ores  which  contain 
the  precious  metals   either   free  or  in   sulphides   or 
chlorides  from  which  the  precious  metals  only  are 
to  be  extracted.    Other  metals  remain  in  the  amalga- 
mation residue. 

REAGENTS:  Cuprous  chloride  dissolved  in  brine;  metals 
which  are  electropositive  to  silver  (Zn,  Pb,  Cu)  used 
in  the  form  of  amalgam;  and  finally  mercury. 
CHEMICAL  REACTIONS:  Cii2Cl2  (or  CuCl)  when 
dissolved  in  a  concentrated  NaCl  solution  will  quickly 
decompose  Ag2S. 

Ag2S  +  2CuCl  *  2  AgCl  +  Cu2S. 

With  the  degree  of  fineness  practically  obtainable 
in  crushing,  not  over  80%  of  the  silver  contained  in 
simple  or  complex  native  sulphides  is  obtained  by 
this  method.  The  stopping  of  the  reaction  cannot 
here  be  attributed  solely  to  the  law  of  mass  action. 
Opposing  the  action  of  the  CuCl  is  the  mechanical 
resistance  offered  by  the  Cu2S  crust  adhering  to  the 
grains  of  Ag2S.  CuCl  alone,  therefore,  is  insufficient 
as  a  solvent  of  Ag2S  in  amalgamation. 

In  the  presence  of  metals  electropositive  to  silver, 
such  as  Zn,  Pb,  and  Cu,  the  Ag2S  is  reduced  to  Ag  in 


42  METALLURGY 

spite  of  the  coating  of  other  sulphides  (Cu2S)  or  salts. 
Cu2S  under  these  circumstances  does  not  hinder 
the  completion  of  the  reaction,  but  acts  as  a  metallic 
conductor  for  the  exchange  of  energy  between  the 
more  positive  metal  and  silver:  it  also  forms  with  these 
and  the  solution  a  galvanic  couple  in  which  Ag2S 
is  the  cathode  and  Zn  is  the  anode.  If  Cu  or  CuCl  is 
lacking  at  the  beginning  of  the  process,  Zn  acts  upon 
Ag2S  as  follows: 


In  spite  of  the  seemingly  complete  separation  of  Ag 
from  Ag2S,  in  this  case  there  is  another  drawback 
which  causes  imperfect  amalgamation.  ZnS,  which 
is  insoluble,  envelopes  the  silver  and  hinders  its 
amalgamation.  By  using  Zn  and  Hg  in  a  pure  NaCl 
solution,  as  much  as  50%  of  the  Ag  that  has  been 
converted  to  the  metallic  state  by  the  Zn  may  remain 
unamalgamated.  Copper  works  more  favorably  and, 
in  fact,  as  Krohnke  pointed  out,  through  the  formation 
of  CuCl,  for  CuCl  causes  the  decomposition  of 
Ag2S,  but  ZnCl2  does  not.  But,  as  stated  above, 
CuCl  alone  is  also  not  sufficient  of  itself  to  completely 
decompose  Ag2S,  and  the  solution  tension  of  Cu  is 
weaker  than  that  of  Zn.  If,  now,  care  is  taken  at 
the  outset  to  provide  for  the  presence  of  CuCl  and 
Zn  is  added,  then  the  following  reaction  takes  place 
without  the  formation  of  ZnS: 

2CuCl  +  Ag2S  +  Zn  =  Ag2  +  Cu2S  +  ZnCl2. 

Cu2S  does  not  hinder  the  amalgamation,  as  it  is  more 
brittle  than  ZnS  and  the  mechanical  movement 
of  the  pulp  easily  exposes  the  surface  of  the  Ag  to 
the  action  of  the  Hg. 


SILVER 


43 


OUTLINE  OF  THE  KROHNKE  PROCESS: 

I.  CRUSHING:    wet  edge-mills  called  Chilian  mills  are 

used. 
2.  COLLECTING  THE  SLIME  in  a  system  of  settling-tanks, 


FIG.  64. — Amalgamation  Barrel.     Length,  6  Feet.  Diameter,  5!  Feet.    Made  of 

3-inch  Pitch  Pine  Planks.     Wooden  (formerly  iron)  rods  pass  through  the 

barrel  between  the  heads  in  order  to  prevent  the  charge  from  balling  up. 
Capacity,  6  tons  of  ore  besides  the  reagents. 

each  of  4000  cu.ft.  (100  metric  tons)  capacity,  in 
which  the  water  is  removed  by  filtration  and 
evaporation. 

3.  PREPARATION  OF  THE  AMALGAM:  Zn   is    heated 
with  ten  times  its  weight  of  Hg  under  acid  water; 


44  METALLURGY 

Pb  with  twice  its  weight  of  Hg  without  acid;   then 
filtered. 

Zn  amalgam  contains  14  to  17%  Zn. 
Pb  amalgam  contains  about  45%  Pb. 

The  Zn  content  may  be  increased  to  58%  by  heating 

and    slowly    cooling    a    large  mass,  in   which  case 

filtration  is  unnecessary. 

APPARATUS  :  cast-iron  kettle  or,  for  large  quantities^ 
a  sheet-metal  kettle  (bottom  of  one  piece).  The 
bottom  edges  should  project  beyond  the  source  of 
heat  so  that  any  mercury  which  trickles  out  or 
is  spilled  may  be  recovered. 

4.  CUPROUS     CHLORIDE     SOLUTION:      i     part     by 
weight  of  CuSO4.5H2O +0.5   to    i%  H2SO4  +  6  to 
20   parts  saturated   NaCl  solutions  +  scrap   copper 
(CuSO4  4-  2NaCl  +  Cu  =  Na2SO4  +  2CuCl)  are  heated 
by  steam   for  from  one  half   to   two   hours   (kept 
saturated    with    NaCl)    in    double-walled    wooden 
vats.       The    space    between  the   walls   is  packed 
with  a  mixture  of  tar  and  powdered  lime. 

5.  AMALGAMATION   is   accomplished   in   barrels    (see 
Fig.  64).     The  barrels  are  charged  in  succession  with 
the  following  reagents,  boiling  NaCl  solution,  Zn 
amalgam  (or  Zn  and  Hg  followed  by  the  rotation  of 
the  barrel  to  produce  amalgam),  then  CuCl  solution, 
followed  immediately  by  the  roughly-broken  lumps 
of  air-dried  slime,  which  disseminates  through  the 
mass,  forming  a  thick  paste.     The  barrels  are  kept 
turning  slowly  (four  or  five  revolutions  per  min.)  for 
four  or  five  hours.     Power  required,  8-9  H.P. 

6.  WORKING   UP  THE  AMALGAM:     After  completing 
the  amalgamation,  the  barrels  are  allowed  to  stand 
for  from  one  to  three  hours;   then  they  are  rotated 
for  a  short  time,  at  first  quickly,  then  slowly.    Every- 


SILVER 


45 


thing  is  then  allowed  to  run  out,  the  barrels  are 
washed  out,  and  the  addition  of  water  is  continued 
until  its  volume  amounts  to  eight  times  that  of  the 
slimes.  The  thinned  slimes  are  worked  up  in  set- 
tling pans,  from  which  the  pulp  gradually  overflows 
and  the  amalgam  is  finally  withdrawn  at  the  bottom. 


FIG.  65. — American- Amalgamating  Pan.  FIG.  66. — Amalgam  Filter. 

The  chemical  purification  of  the  amalgam,  particu- 
larly from  copper,  is  accomplished 

Either  by  treatment  with  Ag2S  or  AgCl; 

By  treatment  with  CuCb-NaCl  liquor; 

Or  by  treatment  with  NH3  and  air. 
The  amalgam  is  finally  filtered  (Fig.  66) ,  this  being 
ultimately  hastened  by  a  centrifugal  machine,  and 
then  it  is  distilled  (Figs.  67  and  68). 


46 


METALLURGY 


CASO  OR  CALDRON  PPOCESS: 
REAGENTS:  Cu,  NaCl,  Hg. 
APPARATUS:  copper  pan. 
APPLICABLE:  only  for  the  purer  ores  carrying  native 


FIG.  67. — The  Krohnke  Furnace  for 
Distilling  Amalgam. 


FIG.  68. — Krupp  Furnace  for 
Distilling  Amalgam. 


silver  or  AgCl.     These  are  seldom  available  and  the 
process  is  little  used  at  present. 
WASHOE  PROCESS: 

REAGENTS:  CuSO4,  NaCl,  Fe,  and  Hg.  The  princi- 
pal reagent,  as  in  the  Krohnke  process,  is  CuCl,  the 
product  obtained  by  the  reaction  mentioned  above, 
under  4.  The  action  of  the  reagents  is  hastened  by 
grinding  between  iron  plates  in  iron  pans  which  are 
heated  by  steam. 


SILVER  47 

AMALGAMATION    APPARATUS:    wooden   or  iron  vats 
with  removable  bottom  plates  and  mullers  (Fig.  65). 
PROCEDURE:  the  ore  is    crushed    in   stamp   mills, 
partially  de-watered  in  settling  tanks   and  then   sub- 
jected to  wet  grinding,  with   the  above  reagents,  in 
steam-heated  pans.     The  pulp  is  subsequently  concen- 
trated and  the  amalgam  purified  and  distilled. 
After  a   preliminary  chloridizing    roast,  ores    may    be 

treated  by 

Barrel  amalgamation. 
Pan  or  vat  (Tina)  amalgamation. 
The  preliminary  treatment  is  essentially  the  same  for 
both  processes:  Preliminary  crushing  (rock  break- 
ers); drying  (usually  in  rotary  cylinders);  further 
crushing  with  salt  (NaCl)  in  stamp  mills  (dry 
stamping)  or  in  other  crushing  apparatus  (edge 
mills  or  ball  mills).  Chloridizing  roast  in  rever- 
beratory  furnaces  (with  stationary  hearths  and 
movable  stirrers — Pearce,  Brown,  and  Wethey;  with 
movable  hearths — Bruckner  and  White-Howell)  or 
in  shaft  furnaces  (Stetef eldt) . 

The  amalgamation  proper  now  takes  place  either  in 
barrels,  in  which  case  there  is  a  preliminary  treat- 
ment with  iron  and  water  to  reduce  ferric  and  cupric 
compounds  to  ferrous  and  cuprous  compounds  and 
to  reduce  AgCl  to  Ag,  after  which  mercury  is  added; 
or  by  pan  amalgamation,  as  in  the  Washoe  process; 
or  by  vat  amalgamation  in  wooden  vats  with  copper 
bottoms  and  copper  mullers.  In  the  latter  case 
the  copper  acts  as  the  reducing  material  for  cupric 
and  ferric  compounds. 
The  amalgamation  procedure  is  the  same  as  that 

described  above. 
Solution  of  the  Silver  in  Chemical  Solvents. 

Ziervogel    Process:    Applicable  to  matte  rich  in  silver.     The 
matte  is  concentrated  to  about  74-75%  Cu  and  carries  at  the 


48  METALLURGY 

Mansfeld  Works  0.4  to  0.5%  Ag.  With  higher  percentages, 
the  Ag  segregates  in  large  pellets  which  is  unfavorable  for 
good  extraction.  The  process  consists  in  converting  Ag  to 
Ag2SO4  and  the  Cu  to  CuO,  leaching  out  and  precipitating 
the  Ag,  and  smelting  the  CuO  for  Cu.  The  operations 
are: 

1.  Breaking  of  the  matte  by  hand. 

2.  Grinding  in  ball  mills. 

3.  Preliminary  roasting  for  two  days  in  a  roasting  furnace  with 
mechanical  stirrers;  this  converts  the  copper  to  CuSO4. 

4.  Final  roasting  in  a  furnace  of  the  same  type,  which  pro- 
duces CuO  and  Ag2SO4. 

5.  Leaching  out  the  Ag2SO4  with  warm  water. 

6.  Precipitation  of  the  silver  with  granulated  copper  in  small 
stoneware  vats  fitted  with  filter  bottoms. 

7.  Washing,  pressing,  drying,  and  fusing  the  silver. 

8.  Re-roasting  the  leached  oxide  and  further  treatment  again 
as  under  5,  6,  and  7. 

9.  The  CuSO4  resulting  from  the  silver  precipitation  is  con- 
centrated and  crystallized  to  form  blue  vitriol  (CuSO4.5H2O). 

In  the  following  processes,  the  Ag  is  either  already  present 
as  AgCl  in  the  ore  or  it  is  converted  into  AgCl  by  a  chloridizing 
roast  and  leached  with  chloride  or  hyposulphite  solution  from 
which  it  is  subsequently  precipitated.     Among  these  processes, 
the  following  have  found  limited  usage : 
The  Old  Oker-Longmaid-Henderson  Process :    leaching  of  the 
pyrite,  after  a  chloridizing  roast,  with  water  from  the  con- 
densation tower  of  the  roasters.     Precipitation  of  the  Ag 
with  Nal  and  the  Cu  with  Fe.    Recovery  of  the  I  by  decom- 
posing the  Agl  with  Na2S. 

Augustin  Process :  the  ore  is  leached  with  a  saturated  brine. 
One  gallon  of  saturated  NaCl  solution  dissolves  at   2o°C. 
0.169  oz-  Ag  (IO°  oz-  NaCl:   0.4  oz.   AgCl).     The  Ag  is 
precipitated  from  the  brine  by  means  of  Cu. 
Kiss  Process:    the  ore  is  leached  with  CaS2Os  solution  and 
the  silver  precipitated  by  Ca(SH)2. 


SILVER  49 

Russell  Process:  leaching  with  4Na2S2O3 -301128203  solu- 
tion and  precipitation  with  Na2S. 

Patera  Process:  this  process  is  used  extensively,  especially 
in  plants  designed  and  built  by  Ottaker  Hofmann.  The 
Patera-Hofmann  process  consists  of  the  following  opera- 
tions : 

1.  A  PRELIMINARY  LEACHING  with  water  to  remove  chlo- 
rides of  the  base  metals. 

APPARATUS:  wooden  vats,  15-20  ft.  in  diameter  and  5  ft. 
deep  with  removable  filter  bottoms  of  lattice  work. 
All  wooden  surfaces  are  coated  with  asphaltum. 

2.  LEACHING  OUT  THE  SILVER  with  sodium  "hyposulphite" 
containing     0.25-0.5%     Na2S2O3-5H2O.  — 100    g.     of 
Na2S2O3-5H2O   will   dissolve  40  g.  of  AgCl,  forming 
Ag2S2O3-Na2S2O3-2H2O.     Same  apparatus  as  in  i. 

3.  PRECIPITATING  of  the  Ag  with  Na2S  or  calcium  poly- 
sulphide,  the  latter  being  prepared  by  boiling  milk  of 
lime  with  sulphur.     The  calcium  polysulphide  solution 
is  preferred    because  it  prevents  the    concentration  of 
Na2SO4  in  the  solutions,  the  SC>4  being  precipitated  as 
CaSC>4.     The  Na2S  resulting  is  oxidized  to  Na2S2O3,  so 
that  no  fresh  supply  of  this  salt  is  required. 

4.  FILTRATION  :  the  precipitate  is  transferred  to  a  settling 
tank  and  from  there  to  a  filter  press. 

5.  DRYING  AND  ROASTING  in  a  reverberatory  furnace. 

6.  CHARGING  INTO  A  HOT  LEAD   BATH  and  cupelling  the 
rich  argentiferous  lead. 

Nitric  Acid  Parting:  This  process,  known  as  inquartation, 
is  only  used  to-day,  except  for  Au-Ag  parting  in  assaying, 
according  as  the  demand  for  the  resulting  AgNOs  permits. 
Ag  is  dissolved  by  HNO3,  leaving  the  Au  unattacked. 
If  the  ratio  of  Au  to  Ag  in  the  alloy  is  between  i :  3  (hence 
the  name  inquartation)  and  1:1.75,  the  Au  remains  behind 
in  a  coherent  state  after  heating  the  alloy  with  HNOs. 
If  the  Au  content  is  higher,  the  resulting  Au  carries  Ag  even 
after  repeated  treatment  with  HNC>3.  With  less  Au, 


50 


METALLURGY 

T 


Plant  for  Nitric  Acid 
Treatment  of  Anode 
Mud  from  Electrolytic 
Parting  of  Gold  and 
Silver.  Globe  Works, 
Denver. 


FIG.  70. — Cross-section  of  Fig.  69. 


SILVER  51 

the  Au  remains  behind  as  a  powder.  Apparatus  necessary : 
glass,  porcelain  or  stoneware  vessels,  less  frequently 
platinum  dishes.  An  apparatus  for  parting  Au-bearing 
anode  mud  by  the  HNOs  method  is  shown  in  Figs.  69-70. 
Sulphuric  Acid  Parting:  applicable  for  Au-Ag  alloys  which 
contain  not  more  than  i  oz.  Au  to  2-4  oz.  Ag  and  less  than 
10%  Cu.  The  H2SO4  dissolves  the  Ag  and  Cu,  forming 
sulphates : 

Ag2  +  2H2SO4  =  Ag2SO4  +  2H2O  +  SO2, 

if  the  granulated  alloy  is  boiled  with  an  excess  of  H2SO4.  On 
a  small  scale,  porcelain  vessels  are  used,  on  a  large  scale, 
cast  iron.  The  acid  solution  is  diluted  with  water  in 
lead-lined  vats  and  clarified  from  small  particles  of  gold 
thereupon  the  silver  is  precipitated  on  scrap  copper  or 
iron  in  other  lead-lined  vats.  The  precipitated  silver 
is  removed  and  compressed.  For  further  treatment, 
see  Silver  Refining.  For  treatment  of  the  gold,  see  Gold 
Refining  (page  21). 

Chloridizing  Fusion:  conducting  Cl  into  the  melted  Au-Ag 
alloy  in  order  to  convert  the  Ag  into  AgCl  is  no  longer  used. 

Electrolysis  of  Ag-Au-Cu  alloys,  see  Copper. 
Solution  of  the  Base  Metals :    In  these  methods  the  silver 
remains  in  the  residue.     According  to  the  nature  of  the  ore 
and  the  condition  of  working,  one  of  the  following  processes 
is  used: 

Freiberg  Vitriolization  Process:  adapted  to  roasted  matte. 
In  contrast  to  the  Ziervogel  process  the  ore  is  so  roasted  that 
the  Cu  is  converted  to  CuO  and  Ag  to  the  metallic  state. 
The  CuO  is  dissolved  as  sulphate  by  treatment  with 
H2SO4.  The  silver-bearing  residue  is  smelted  with  silver- 
lead  ores  and  the  solution  is  worked  up  into  blue  vitrol 
(CuSO4-5H2O). 

Rossler's  Sulphurizing  Fusion.  This  consists  in  melting 
Au-Ag-Cu  alloys  with  S.  There  remains  an  Au-Ag  alloy, 


52  METALLURGY 

Electrolysis  Mobius  Process. 


FIG.  71. — Cross-section  of  Electro- 
lytic Cell. 


FIG.  72. — Anode  Cell  with  Five  Anode 
Plates  of  Dore  Silver. 


FIG.  73. — Cathode  and  Supporting  Device. 


FIG.   74. — Modern   Mobius  Apparatus.       Cathode,  endless   belt  of    sheet  silver 

having  below  devices  for  greasing  and  brushing.  This  belt  is  drawn  through 

a   shallow    vat.       Anodes,  bars  of  dore   silver  placed    in  shallow    troughs 
having  porous  bottoms. 


SILVER  53 

which  is  parted  either  by  the  sulphuric  acid  method  (see 
above)  or  by  electrolysis  (see  Silver  Refining).  The 
Cu2S,  which  contains  some  Ag2S,  is  converted  to  argentifer- 
ous copper  by  an  oxidizing  smelt  and  then  electrolyzed. 
Hartz  Vitriolization  Process.  This  process  converts  precious 
metals  carrying  copper  into,  blue  vitriol  by  the  action 
of  air  and  dilute  sulphuric  acid  on  the  granulated  alloy : 

H2SO4  +  O  +  Cu  =  CuSO4  +  H2O. 

The  solution  is  clarified  and  converted  by  concentration 
into  blue  vitriol.  The  slime  carrying  the  precious  metal, 
after  being  washed  and  dried,  is  refined  by  cupellation. 

Electrolytic  Refining:    see  Copper. 

Electrolytic  Refining  of  Zinc  Skimmings :    see  Zinc. 

(B)   Refining  Silver 

The  aim  in  silver  refining  is  to  remove  the  last  traces  of  impur- 
ities and  to  effect  also  the  separation  of  the  silver  from  other  pre- 
cious metals.  It  consists  usually  of  an  oxidizing  fusion  in  which 
oxidizing  fluxes  are  utilized  as  well  as  the  atmospheric  oxygen, 
and,  if  other  precious  metals  are  present,  of  electrolysis. 

1.  Fire  Refining  of  the  crude  silver  either  by  melting   under 
airblast,  i.e.,   carrying  out  cupellation  in  small  reverberatory 
furnaces  with  oxidizing  flame,  or 

FUSING  WITH  NITER,  which  is  common  in  working  up 
precipitated  silver  (cement  silver)  which  carries  small 
quantities  of  Fe  or  Cu  (apparatus:  crucibles),  or  by 

FUSING  WITH  Ag2SO4  (Rossler's  method),  adapted  to  Ag 
carrying  Bi.  This  process  is  also  carried  out  in  crucibles. 

2.  Electrolytic  Refining.    According  to  Mobius  and  Wohlwill, 
if    gold-bearing    silver    (the    so-called    dore    silver)    in    the 
form    of    small  plates  or  bars,   is  used  as  the  anode  in  an 
aqueous  solution  of  HNO3  (i%),  AgNO3  (0.5%  Ag)  and 
Cu(NO3)2    (4%    Cu),    the  Ag   will   be    dissolved    and   de- 
posited upon  a  sheet  of  pure  Ag,  which  serves  as  the  cathode. 


54 


METALLURGY 


The    current   density    recommended  is  30  to   20    amperes 
per  sq.ft.     The  E.M.F.  is  1.5-1.4  volts.     The  apparatus  is 
shown  in  Figs.  71,  72,  73,  and  74,  page  52. 
Properties  of  Refined  Silver : 
SPECIFIC  GRAVITY:   10.5. 


FIG.  75. 


FIG.  76. — Fused  Silver.     Dendritic  Surface  (X33). 

COLOR:  white  with  brilliant  lustre. 

DUCTILITY:  tough,  very  ductile. 

MELTING-POINT:  961°  C.  (1762°  F.). 

VAPORIZED  readily  at  1200-1500°  C.  (2192-2732°  F.). 


SILVER 


55 


THERMAL  AND  ELECTRICAL  CONDUCTIVITY  best  of  all  metals 
ALLOYS  WITH  MOST  METALS:   the  most  important  metal- 
lurgical alloys  are  those  with  Au,  Pb,  Hg,  Cu  and  Zn. 


•  ^^ 

^&Mim 


FIG.  77. — Electrolytic  Silver  (Mag.  16). 

CHEMICAL  BEHAVIOR:  not  very  active,  although  its 
solution  tension  is  greater  than  that  of  gold.  Silver 
compounds  are  easily  dissociated  and  are  sensitive  to  light 
(photography).  As  solvents  for  metallic  Ag,  concentrated 
HNO3  and  H2SO4  are  used.  For  AgCl  other  metallic 
chlorides  and  "  hyposulphite  "  are  employed. 


MERCURY 

Sources 

Natural  Sources  : 

FREE,  usually  containing  Ag  and  less  frequently  Au. 

As  CHLORIDE,  HgCl,  calomel  (rare). 

As   SULPHIDE,  HgS,   cinnabar,   the   most  important  ore  of 

mercury.     It  is  finely  disseminated  through  the  gangue  and 

the  per  cent  of   mercury  present   is    usually  low.      With 

other  sulphides  in  tetrahedrite. 
Other  Sources  : 

AMALGAMS  resulting  from  the  use  of  mercury  in  extracting 

other  metals  or  from  contact  with  metals  in  other  ways. 
RESIDUES  and  unmarketable  products  produced  in  making 

mercury  pigments  (HgO  and  HgS). 

(A)    Extraction  of  Mercury 

Roasting  Accompanied  by  Distillation  : 

Oxidizing  Roast  is  the  method  commonly  used  in  treating 
mercury  ores.  Since  the  dissociation  temperature  of  HgO  is 
very  low  (400°  C.),  Hg  results  directly  from  roasting  at  high 
temperatures  : 


APPARATUS:    Shaft  furnaces  for  lump  or  briquetted  ores. 

(Figs.  78-79.) 
SELF-FEEDING  ROASTERS  for  finely-divided  ores.     (Figs. 

80-83.) 
REVERBERATORY  FURNACES  for  easily  reducible  or  sintering 

ores.     To  all  of  these  furnaces,  a  system  of  iron  or  clay 

56 


MERCURY 


57 


CONDENSATION    TUBES    is    attached,    these    tubes    being 

partly  cooled  by  air,  and  partly  by  water. 
The  products  of  these  processes  are: 

LIQUID  MERCURY,  which  collects  under  .water  in  the  lowest 
part  of  the  condenser. 

MERCURIAL  DUST  or  SOOT,  made  up  of  flue  dust  from  the 


FIG.  79. 

ore  and  sublimated  mercury  compounds  (HgO,  HgS, 
HgSO4)  which  collect  on  the  walls  of  the  condenser 
and  mechanically  retain  globules  of  mercury. 


58 


METALLURGY 


MERCURY  59 

Heating  with  Desulphurizing  Fluxes  (Fe  and  CaO),  is 
used  only  for  especially  rich  ores,  or  for  residues  from  the 
manufacture  of  cinnabar.  For  this  purpose  retorts  are  used 
such  as  have  been  already  described  under  Amalgam  Dis- 
tillation (see  Silver). 

(B)   Mercury  Refining 

Filtration  of  the  liquid  mercury  that  collects  in  the  condensing 
tubes  is  frequently  sufficient.  The  use  of  presses  is  necessary 
in  working  up  the  "  dust  "  or  "  soot." 

Distillation  is  used  for  amalgam  (see  Silver). 

Washing  with    Acids    is  employed  when  the  mercury  is  to  be 

used  for  scientific  purposes. 
Properties  of  Mercury : 

SPECIFIC  GRAVITY:    13.5. 

COLOR  :    bluish- white. 

TENACITY:    liquid  at  ordinary  temperatures. 

FREEZING-POINT:    -39.4°  C.  (-38.9°  F.). 

VOLATILIZES  slowly  at  ordinary  temperatures. 

BOILING-POINT:  360°  C.  (680°  F). 

ELECTRICAL  CONDUCTIVITY:    about  0.017  tnat  °f  Ag. 

ALLOYS:  form  readily  with  Au,  Ag,  Cu,  Pb,  Sn,  Cd,  Zn,  with 
the  alkalies  and  alkaline  earths.  More  difficultly  with  Ni, 
Co,  Fe,  Mn,  Cr,  W  and  the  earths. 

CHEMICAL  BEHAVIOR:  The  affinity  of  Hg  for  the  halogens 
and  for  S  is  great;  slight  for  oxygen.  The  chlorides  HgCl 
and  HgCl2,  as  well  as  the  sulphide  HgS,  may  be  sublimed 
without  decomposition.  HgO  first  forms  at  about  300°  C., 
but  dissociates  very  readily  at  400°.  The  best  solvents  for 
Hg  are  HNO3  and  aqua  regia.  Hg  may  form  a  univalent 
as  well  as  a  bivalent  cation  in  solution. 


COPPER 

Sources 

Natural  Sources: 

NATIVE  COPPER:  surrounded  or  accompanied  by  sulphide 
or  oxide  copper  ores,  clay  and  shale. 

RED  COPPER  ORE,  cuprite,  Cu2O,  carrying  88.8%  Cu.  Asso- 
ciated with  sulphide  copper  ores,  spathic  iron  ore,  clay 
or  shale,  and  earth. 

BLACK  COPPER  ORE,  thenorite,  CuO  (rare). 

COPPER  GLANCE,  Cu2S,  free  but  not  widely  distributed. 

COPPER  PYRITES,  chalcopyrite,  Cu2S.FeS.FeS2.  This,  the 
most  important  copper  ore,  contains  34.6%  Cu.  It  is 
associated  with  galena,  blende,  shale  and  slate. 

PEACOCK  ORE,  bornite,  (Cu2S)3.FeS.FeS2.  Not  so  abun- 
dant as  chalcopyrite. 

BLUE  VITRIOL,  CuSO4>5H2O.  Formed  by  the  weathering 
of  sulphide  ores  and  associated  with  their  gangue. 

MALACHITE,  or  green  carbonate  of  copper;  HO-Cu-COs-Cu- 
OH;  formed  by  the  weathering  of  other  copper  minerals. 
Associated  with  similar  minerals  and  in  similar  localities. 

AZURITE,  or  blue  carbonate  of  copper;  HO-Cu-COs-Cu-COy- 

Cu-OH;  associated  with  malachite. 
Other  Sources: 

MATTE  containing  copper,  speiss,  slag  and  alloys  from  lead 
and  nickel  smelting. 

SCRAP  METAL  and  residues  from  metal-working  processes. 

(A)   Enriching  and   Concentrating  Processes 

The  percentages  of  Cu,  as  given  above,  refer  to  the  pure  mineral, 
free  from  gangue.  By  means  of  the  gangue,  however,  the  Cu 
content  of  most  ore  bodies  is  reduced  to  a  few  per  cent.  From 

60 


COPPER 


61 


these  ores  the  copper  can 
neither  be  successfully 
leached  out  by  chemical 
means  nor  can  it  be  suf- 
ficiently concentrated  by  me- 
chanical processes  so  that 
the  ores  can  be  successfully 
fused  directly  for  metal,  as 
this  would  be  so  impure 
that  refining  would  be  out  of 
the  question.  It  is  neces- 
sary, therefore,  to  effect  a 
chemical  concentration  by  a 
fusion.  Of  all  the  metals 
that  come  into  considera- 
tion here,  copper  has  the 
greatest  solution  tension  in 
fused  sulphides;  it  will, 
therefore,  withdraw  the 
sulphur  from  other  metals  in 
so  far  as  the  total  amount  of 
sulphur  present  in  a  fused 
mixture  of  sulphides  and 
other  compounds  is 
insufficient  to  form  the 
lower  sulphides  of  all 
the  metals.  Upon  this 
principle  are  based  the 
following 
Roasting  and  Smelting 

Operations : 

i.  An  Oxidizing  Roast 
effects  the  removal  of 
the  sulphur  from  sul- 
phide ores,  leaving, 
however,  more  than 


62 


METALLURGY 


enough,  after  allowing  for  a  loss  due  to  reaction  in  the  furnace, 
to  form  Cu2Swith  the  copper  present.   In  matte  smelting,  there 


SECTION  A-B 


FIG.  85.— Stall  Roasting  Plant  (Peters).     Scale,  i :  150. 

must  always  be    enough    sulphur    present    to    form 

with  the  copper  and  to  form  a  certain  amount  of  FeS,  to 

prevent  a  high  loss  of  copper  in  the  slag;  the  FeS  and  Cu2S 


COPPER  63 

should  be  present  in  such  quantities  as  to  correspond  to  the 
formula  FeS.Cu2S. 

In  crushing  the  ore,  which  is  usually  in  the  form  of  large 
lumps  to  make  it  suitable  for  roasting,  it  should  be  observed 
that  the  finer  the  ore  the  greater  the  cost  of  roasting,  the 
greater  the  loss  in  flue-dust,  and  the  greater  the  hindrance 
in  subsequent  leaching  operations.  The  size  of  grain 
desirable  in  roasting  is  from  6  to  60  mm.  In  choosing  a 
method  of  crushing,  one  that  is  quick  and  gives  the  smallest 
amount  of  fines  should  be  selected.  Hand  breaking  has 
advantages,  for  the  fines  are  then  under  10%  as  a  rule. 
Machines  which  produce  the  smallest  quantity  of  fines 
usually  require  the  most  power.  If  the  nature  of  the  ore 
requires  that  it  be  reduced  to  a  small  grain,  rolls,  stamp 
mills  and  ball  mills  are  used.  The  following  processes  and 
and  apparatus  may  be  mentioned. 

HEAP  ROASTING:  Open  heaps  of  ore  having  a  trapezoidal 
or  hemispherical  cross-section  are  constructed  upon  some 
combustible  material  in  order  to  start  the  roasting.  (See 
Fig.  84.) 

STALL  ROASTING  :  To  serve  as  a  protection  against  the  weather 
and  to  prevent  damage  to  the  neighborhood,  walls  may  be 
built  to  protect  the  heap.  These  walls  at  first  consisted 
of  board  fences,  earth  banks,  and  finally  stone  walls, 
from  which  came  the  so-called  "  stall."  Stall  roasting  is 
essentially  the  same  as  heap  roasting  except  for  the 
protective  walls,  which  also  serve  to  regulate  the  amount 
of  air.  The  best  stall  construction  is  Peters'  modifica- 
tion of  the  Swedish  stalls.  (See  Fig.  85.) 
PYRITE  BURNERS:  these  may  be  regarded  as  a  further  step 
in  stall  construction.  Many  different  forms  have  been 
developed,  among  which  may  be  mentioned:  kilns  (Fig- 
86),  lump  pyrites  burner  (Fig.  87),  fine  pyrites  burner, 
Gerstenhofer's  fine  pyrites  furnace,  Hasenclever-Helbig 
furnace,  Olivier-Ferret  furnace  and  the  Maletra-Shaffner 
roasting  furnace  (Fig.  88). 


64 


METALLURGY 


COPPER 


65 


REVERBERATORY     ROASTING.     As     compared    with    other 
reverberating  furnaces,  the  following  points  are  character- 


Fic.  89. — Stetefeldt  Shaft  Furnace. 

istic:  They  should  combine  a  small  firebox  with  a  rela- 
tively long  hearth,  and  the  length  of  the  latter  depends  upon 
the  amount  of  heat  evolved  from  the  roasting  ore.  Ores 


66 


METALLURGY 


carrying  10%  or  less  of  sulphur  are  seldom  roasted  in 
reverberatories.  The  relation  between  length  of  hearth 
and  per  cent  of  sulphur  is  as  follows: 

With  10%  sulphur,  length  of  hearth  15  ft. 
With  15%  sulphur,  length  of  hearth  30  ft. 
With  20%  sulphur,  length  of  hearth  45  ft. 
With  25%  sulphur,  length  of  hearth  60  ft. 


FIG.  90. — MacDougall  Furnace.  FIG.  91. — Herreshoff  Furnace. 

Hearth-furnaces  worked  by  hand  are  designated  as  hand 

reverberatories. 
Among  the  SHAFT  FURNACES  the  Stetefeldt  furnace  should 

be  mentioned.     It  is  especially  suitable  for  chloridizing 

roasting.    (Fig.  89). 
The  most  important  reverberatories  with  mechanical  stirrers 

are  the  following: 
PARKES    FURNACE:    this    has    two    circular    superimposed 

hearths   with   radial  stirring  arms  attached   to  a  central 

shaft.    The  firebox  is  beside  the  lower  hearth. 


COPPER  67 

MACDOUGALL  FURNACE  (Fig.  90),  Herreshoff  furnace  (Fig. 
91)  and  the  furnaces  of  Humboldt  and  Kaufmann:  these 
have  five  to  seven  circular  hearths  placed  one  above  the 
other.  Horizontal  arms,  extending  from  a  vertical  shaft,, 
work  the  ore  in  such  a  way  that  on  the  upper  hearth  it  is 
slowly  moved  toward  the  periphery.  Here  it  falls  to  the 
next  hearth,  is  worked  toward  the  centre,  falls  to  the  third: 
hearth  and  so  on  to  the  last.  The  arms  are  easily  removed 
and  replaced  by  new  ones. 

O'HARA  FURNACE:  In  its  original  construction  this  furnace 
consisted  of  two  superimposed  hearths  through  which 
rakes  resembling  ploughshares  were  drawn  by  means  of  an 
endless  chain.  The  latter  was  run  upon  sprocket  wheels. 
During  interruptions  in  the  process,  the  chains  should  be 
dropped  into  a  groove  in  the  hearth  in  order  to  protect 
them  from  the  hot  gases.  This  apparatus  is  untrustworthy 
and  suffers  greatly  from  hot  gases. 

ALLEN  FURNACE:  Allen  obviated  the  above  difficulties  by 
placing  tracks  in  his  furnace  upon  which  carriages  carry- 
ing the  rabbles  were  supported.  The  carriages  were 
drawn  by  the  chain  as  before. 

BROWN  FURNACE:  Brown  placed  the  tracks  and  supporting 
carriages  in  compartments  built  in  the  side  walls  of  the 
furnace.  Brown's  furnace  is  also  constructed  in  the  form 
of  a  broken  circle,  the  so-called  elliptical  furnace,  with 
external  fireboxes.  (Fig.  92.) 

HIXON  FURNACE:  Hixon  finally  placed  the  supporting 
trucks  and  the  dragging  device  (iron  cables  instead  of 
chains)  in  a  channel  built  in  the  bottom  of  the  furnace. 

HOLTHOFF-WETHEY  FURNACE  :  the  rails  and  trucks  for  the 
tube-like  rabble  holders  were  placed  entirely  outside  the 
walls  of  the  furnace.  The  space  beneath  the  hearth  served 
mainly  as  a  cooling  hearth. 

PEARCE  FURNACE  (see  Gold,  page  3,  Figs,  i  and  2): 
The  hearth  is  circular.  The  rabbles  are  held  by  tube-like 
arms  through  which  air  for  roasting  is  supplied.  The 


68 


METALLURGY 


FIG.  92. — O'Hara-Brown  Elliptical  Furnace. 


COPPER 


69 


rabble  arms  are  driven  from  a  central  shaft  and  project 
through  a  slot  in  the  inner  wall  of  the  furnace.  The  slot 
is  kept  closed  by  an  iron  ring  which  moves  with  the  rabbles. 
The  Pearce  furnace  is  built  both  with  a  single  hearth  and 
with  two  hearths  one  above  the  other.  Auxiliary  heating, 
if  necessary,  is  supplied  by  external  fire  boxes.  The  space 
under  the  roasting  hearth  serves  as  a  dust  chamber. 
WETHEY  FURNACE:  This  is  a  furnace  for  fine  pyrite  and 
consists  of  four  superimposed  hearths.  The  rabbles  make 
a  circuit  through  two  hearths.  Two  of  these  furnaces 


{- -3658- 


FIG.  93. — Bruckner  Furnace.      Scale   1:150. 

may  be  placed  side  by  side  and  the  rabbles  moved  by  the 
same,  machinery. 

SPENCE  FURNACE:  This  is  a  furnace  for  treating  fine  pyrite. 
It  consists  of  four  superimposed  hearths,  the  rabbles  being 
drawn  back  and  forth  over  the  hearths  by  means  of  horizon- 
tal rods. 

KELLER  FURNACE:  This  is  a  furnace  for  fine  pyrite.  It 
has  five  superimposed  hearths  and  its  operation  is  similar 
to  that  of  the  Spence  furnace.  Two  furnaces  are  usually 
placed  side  by  side  and  operated  by  the  same  engine. 

Among  furnaces  with  moving  hearths  the  following  have 
found  favor. 


70  METALLURGY 

BRUCKNER  REVOLVING  FURNACE  (Fig.  93) :  This  furnace, 
which  is  widely  used,  consists  of  a  cylinder  resting  upon 
friction  wheels  to  which  power  is  applied  by  means  of 
toothed  gears.  The  heated  gases  from  an  external  fireplace 
pass  axially  through  the  cylinder.  At  the  opposite  end  are 
the  dust  chamber  and  stack.  The  cylinder  is  charged 
from  hoppers  placed  above  manholes  in  the  furnace 
walls,  the  holes  being  subsequently  closed.  After  the 
roasting  is  completed  the  furnace  is  discharged  through 
the  same  holes  into  cars. 

WHITE-HOWELL  FURNACE  (Fig.  94):  The  roasting  hearth 
consists  of  a  slightly-inclined  cylinder,  open  at  both  ends. 


FIG.  94. — White-Howe  11  Furnace. 

It  is  made  of  sheet  iron  and  lined  with  firebrick.  The 
furnace  is  fed  through  a  tube  at  the  higher  end  and  is 
discharged  into  a  walled  chamber  at  the  lower  end.  Beside 
this  chamber  is  the  fireplace  from  which  the  hot  gases  pass 
over  the  roasting  ore.  At  the  higher  end  is  a  dust  chamber 
from  which  leads  the  flue  to  the  stack. 

HOCKING- OXLAND  FURNACE  :  This  is  similar  in  construction 
to  the  White-Howell,  but  somewhat  differently  built  in 
the  arrangement  of  the  firebox  and  the  chamber  for  collect- 
ing the  roasted  ore. 

ARGALL  FURNACE  (Fig.  95):  Within  a  broad  cylinder  are 
enclosed  four  smaller,  somewhat  shorter  cylinders  slightly 
inclined.  The  large  cylinder  is  fitted  with  firebox  and 


COPPER  71 

dust  chamber  similar  to  those  previously  described.  The 
ore  is  fed  into  each  narrow  cylinder  in  turn  as  it  reaches 
its  highest  point  during  the  revolution  of  the  broad  cylinder, 
and  is  discharged  at  the  opposite  end,  which  lies  next  to 
the  firebox.  A  chamber  below  the  large  cylinder  receives 
the  ore. 

MUFFLE  FURNACES:  These  are  used  for  the  chloridizing 
roasting  of  copper  ores,  especially  roasted  pryite  carrying 
copper.  The  construction,  shown  in  Fig.  96,  page  72,  has 
four  iron  charging  tubes  with  gates  for  closing.  The  charge 
is  worked  by  hand  through  doors  in  the  side  of  the  muffle. 


FIG.  95. — Argall  Furnace.     Scale,  i  :  150. 

The  gases  are  conducted  away  through  an  iron  tube  near 

the  firebox.      In  order  to  lessen  the  danger  of  overheating 

the  charge,  the  muffle  arch  is  made  of  double  thickness. 

2.  Smelting  for  Matte  consists  in  uniting  the  copper  of  the  roasted 

product  with  part  of  the  sulphur,  uniting  the  rest  of  the  sulphur 

with  part  of  the  iron  and  slagging  the  remaining  iron  together 

with  other  materials.     In  order  to  prevent  high  slag  losses, 

it  is    best  not    to    attempt    too    great    a    concentration  of 

copper.     A  matte  carrying  50%  copper  should  be  regarded 

as  the  maximum  and  in  most  cases  a  lower  grade  is  produced. 

The  concentration  of  the  ore  to  matte  obtained  by  these 

processes  fluctuates  widely,  viz.,   between  12:1  and  3:1  or, 

in  other  words,  between  12  and  3  tons   of  ore  yield    i   ton 


72 


METALLURGY 


COPPER  73 

of  matte  by  the  process  of  roasting  and  smelting.    In  roasting 

attention  must  be  paid  to  these  limits. 

The  roasted  product  contains  a  mixture  of  the  oxides,  sul- 
phides, sulphates,  arsenates,  antimonates,  and  silicates  of 
the  various  metals  contained  in  the  ore. 

PRINCIPLES  OF  THE  SLAG  CALCULATION:  The  slag  should 
lie  between  a  singulo-  and  a  bi-silicate,  preferably  in  the 
vicinity  of  a  singulo-silicate  with  FeO  as  the  predominating 
base.  Exceptions:  With  high  iron  and  zinc  content: 
between  a  singulo  and  a  sub-silicate.  With  high  SiC>2 
content :  between  a  bi-silicate  and  a  tri-silicate.  It  should 
be  borne  in  mind  that  the  furnace  temperature  must  be 
raised  if  high  silica  is  used,  thus  increasing  the  danger 
of  reducing  the  iron  compounds : 

FeS  +  Cu2  =  Cu2S+Fe. 

APPARATUS  FOR  MATTE  SMELTING: 

i.  SHAFT  FURNACES:  From  the  original  narrow  shafts 
with  rectangular  or  trapezoidal  cross-sections,  a  change 
has  been  made  in  favor  of  wider  furnaces  with  circular, 
oval,  or  long  rectangular  cross-section,  and  recently  to  six-, 
seven-  and  eight-sided  sections. 

The  height  of  the  shaft  depends  upon  the  iron  content  of 
the  ore  and  the  nature  of  the  fuel.  The  height  diminishes 
as  the  iron  content  increases.  Furnaces  using  charcoal 
should  be  made  higher  than  those  using  coke.  High  zinc 
necessitates  a  low  furnace,  as  otherwise  deposits  of  zinc 
oxide  will  be  formed. 

The  chemical  action  of  the  matte  and  slag  have  led  to  the 
adoption  of  cooled  furnace  walls.  Whereas  in  lead  smelting 
only  a  relatively  small  portion  of  the  wall  is  cooled,  in 
copper  smelting  not  only  the  whole  bosh  but  in  some  cases 
the  entire  shaft  is  surrounded  by  a  water-jacket.  For  these 
relatively  large  cooling  plates,  cast  iron  should  be  avoided 
as  much  as  possible.  Besides  wrought  iron,  copper  is 


74 


METALLURGY 


FIG.  97.— Early  Type  of  Shaft  Furnace. 


COPPER 


FIG.  98.— Early  Type  of  Shaft  Furnace. 
Scale,  i  :  150. 


FIG.  99. — Oker  Furnace  in  the 
Harz. 


76 


METALLURGY 


sometimes  used,  especially  for  the  inside  walls.  To 
lessen  the  danger  of  interrupting  the  process  by  an 
accident  to  the  water-jacket,  the  latter  should  never  be 
constructed  of  one  piece,  but  in  segments  which  may  be 
independently  removed  and  replaced. 

The  first  requisite  for  the  use  of  a  water-jacket  is  an 


FIG.  100. — Allis-Chalmers.         FIG.  101. — Colorado  Iron  Works,  Denver. 
Scale,  i  :  150. 

adequate  supply  of  water.  If  this  is  available,  there  is  no 
danger  connected  with  the  process. 

In  construction,  operation  and  repairing,  the  water- 
jacket  furnace  is  to  be  preferred  to  a  brick  furnace. 

With  good  management  the  cost  of  operation  is  about 
the  same  for  each.  Great  care  should  be  taken  with  brick 
furnaces  in  the  choice  of  bricks,  and  of  mortar,  and  in  the 
construction.  Water-jacket  furnaces  are  easier  to  design, 
also  easier  and  quicker  to  build. 

Brick  furnaces  also  require  great  care  and  attention 
when  blowing  in.  If  the  blast  pressure  is  too  high,  great 
difficulty  may  be  encountered,  such  as  burning  out  the 


COPPER 


77 


hearth  and  shaft  walls.  If  the  blast  pressure  is  too  low, 
wall-accretions  and  sows  may  form.  Scarcely  any  difficulty 
is  experienced  in  blowing  in  a  water- jacketed  furnace. 

With  similar  handling,  the  relation  between  the  repair 
costs  of  brick  and  water-jacket  furnaces  is  2:1.     Cracks 


FIG.  102. — American  Water-jacket  Furnace  (20  Tuyeres),  Colorado  Iron  Works. 

Scale,  i  :  150. 

and  thin  places  formed  in  the  walls  of  brick  furnaces 
tend  to  grow  worse  instead  of  better,  for  even  if  they  are 
filled  with  clay,  the  matte  and  slag  tend  to  eat  under  the 
clay  and  widen  the  crack. 

If  in  order  to  remove  accretions  it  is  necessary  to  lower 
the  charge,  it  is   often   found   in    brick  furnaces  that  the 


78 


METALLURGY 


FIG.  103. — Sticht  Shaft  Furnace,  Mount  Lyell  M.  &  R.  Co.,  Tasmania.     Side 
View  and  Longitudinal  Section. 


COPPER 


79 


accretions  stick  even  more  firmly  to  the  walls  than  before, 
.because  of  the  fact  that  the  upper  part  of  the  walls  of  the 


FIG.  104.— Sticht  Shaft  Furnace,  Mount  Lyell  M.  &  R.  Co.,  Tasmania.     End 
View  and  Cross-section. 

empty  shaft  have  become  very  hot.  In  the  water-jacketed 
furnace,  however,  the  accretions  can  be  readily  removed, 
for  the  walls  remain  cool. 


80 


METALLURGY 


FIG.  105. — Johnson  American  Water- 
jacket  Furnace  (16  Tuyeres).  Side 
Elevation. 


FIG.  106. — Johnson  American    Water- 
jacket  Furnace  (16  Tuyeres).     Side 

Elevation. 


COPPER  81 

SLAG  AND  MATTE  SPOUTS  :  In  small  furnaces  these  are 
made  of  brasque,  in  large  furnaces  of  cooled  plates  formed 
by  fastening  wrought-iron  tubes,  about  one  inch  in  diameter, 
to  cast-iron  troughs.  (Hixon's  method  of  building  Ltir- 
mann's  slag  spout.) 

FORE-HEARTHS.  To  lessen  undesirable  reactions  which 
result  in  the  formation  of  furnace  "  sows,"  copper  losses 
and  other  difficulties,  it  is  of  great  importance  in 
copper  blast-furnace  smelting  that  the  matte  be  removed 
as  soon  as  possible  from  the  furnace.  The  furnaces  are 
therefore  almost  always  built  with  external  crucibles 
so  that  the  separation  of  matte  and  slag  may  take  place 
in  the  fore-hearth. 

The  first  fore-hearths  were  made  of  brasque  in  the  form 
of  holes  in  front  of  the  furnace.  (See  Figs.  97  and  98, 

PP-  74,  750 

The  modern  fore-hearths  consist  either  of  portable  iron 
pots  or  larger  boxes  (Figs.  104  and  106,  pp.  79,  80),  or  of 
special  reverberatory  furnaces  for  acid  slags.  The  smaller 
pots  are  lined  with  clay  and  straw,  the  larger  ones 
are  lined  a  half  course  of  firebrick.  If  the  fore-hearth  is 
very  large  the  lining  is  omitted  entirely.  The  first  slag 
chills  on  the  sides  of  the  hearth  and  the  heat  radiated 
from  the  walls  so  regulates  the  temperature  that  the  slag 
remains  as  a  lining.  A  diameter  of  10  ft.  is  the  upper 
limit  for  brick-lined  hearths. 

Watercooling  is  resorted  to  in  some  cases  for  the  whole  wall 
of  the  hearth  as  well  as  for  special  parts,  for  example  the  matte 
tap,  the  slag  spout,  and  in  the  so-called  Orford  fore-hearth 
for  the  partition  which  separates  the  matte  from  the  slag. 

According  to  Peters  the  use  of  reverberatories  as  fore- 
hearths  possesses  the  following  advantages: 

(a)  Economy  in  the  cost  of  remelting  for  subsequent  treat- 
ment of  the  matte. 

(b)  Reduction  of  the   necessary  temperature  required  in 
the  blast  furnace  and  thus  saving  of  fuel. 


METALLURGY 

(c)  Prevention  of  overdriving  the  blast  furnace,  for  if  little 
matte  is  required  at  any  period  of  the  subsequent  opera- 
tions, the  time  can  be  utilized  in  accumulating  a  small 
stock  of  especially  rich  or  poor  matte,  casting  it  into 
bars  of  definite  size  and  laying  it  to  one  side.    A  means  is 
thus  provided  for  regulating  the  amount  and  composition 
of  subsequent  charges  in  the  reverberatory. 

(d)  Assistance  in  the  separation  of  matte  and  slag.     In 
the    unheated    fore-hearth,    the    temperature    is    con- 
tinually falling,  while  in  the  reverberatory  fore-hearth 
the  products  of  the  blast  furnace  are  run  into  a  heated 
chamber,  which  favors  the  separation    of    matte   and 
slag. 

The  reverberatories  used  as  fore-hearths  do  not  differ 
essentially  from  other  copper  reverberatories.  The 
following  features,  however,  should  be  provided: 

(a)  The  matte  from  the  blast  furnace  should  be  introduced 
at  one  of  the  ends. 

(b)  The  slag  should  be  withdrawn  from  the  opposite  end. 

(c)  The  fireboxes  are,  therefore,  on  the  sides  and  should 
be  arranged  so  as  to  close  easily. 

(d)  The  molten  matte  should  flow  as  directly  as  possible 
from  the  fore-hearth  to  the  converters. 

(e)  Sufficient  space  should  be  provided  to  allow  50  tons 
of  matte  to  be  withdrawn  and  stored  in  case  of  an  inter- 
ruption in  the  subsequent  processes. 

The  matte  is  withdrawn  from  the  reverberatories 
either  into  a  sand  bed  where  it  cools  and  is  broken 
into  lumps,  or  else  it  is  run  into  casting  ladles  for  delivery 
in  the  molten  state  to  the  next  apparatus. 

When  a  reverberatory  fore-hearth  is  not  used,  the  slag 
is  allowed  to  flow  through  small  fore-hearths  (slag  pots) 
in  which  a  small  quantity  of  matte  settles.  From  these 
slag  pots  it  is  withdrawn  to  be  made  into  building  stone 
if  sufficiently  acid,  or  it  flows  into  slag  cars  for  transpor- 
tation to  the  dump.  If  the  location  and  water  supply  is 


COPPER 


83 


,84  METALLURGY 

favorable,  it  is  granulated  in  a  stream  of  water  and  by 
it  sluiced  to  the  dump. 

In  smelting  for  matte,  and  also  in  the  process  of  matte 
concentration  to  be  described  under  (4),  there  is  oppor- 
tunity to  add  any  oxidized  ore  and  smelter  products 
(slag  rich  in  copper)  that  are  at  hand,  and  work  them 
up  at  the  same  time,  but  care  must  be  taken  that  in  the 
roasting  or  in  the  mixing  of  the  charge  enough  sulphur 
is  present,  or  made  up  by  the  addition  of  unroasted 
sulphide  ores,  to  provide  enough  sulphur  for  all  the  Cu 
that  is  in  the  form  of  oxide. 

2.  REVERBERATORY  FURNACES  FOR  MATTE  SMELTING  are 
used  at  several  American  smelting  plants  on  a  very  large 
scale.  The  reverberatory  process  is  used  principally  for 
finely  divided  ores  and  roasted  products. 

According  to  Peters  (Metallurgie,  2,  9,  1905),  the  grate, 
hearth,  flues  and  stack  must  be  so  designed  that  a  tem- 
perature of  1700°  C.  can  be  developed  if  from  1400°  to  1500° 
is  required  to  fuse  the  charge.  The  firebox  and  hearth 
must  be  of  such  size  that  fresh  additions  of  coal  or  ore  will 
not  cause  a  great  fall  in  the  temperature.  The  hearth 
made  of  fused  sand  must  be  kept  covered,  to  a  depth  of  at 
least  8  in.,  with  fused  matte  to  resist  the  action  of  the 
fresh  charge  which  bakes  on  to  the  sides  and  of  the 
oxides  that  form  during  the  operation.  The  hearth 
should  be  of  such  length  as  to  best  utilize  the  heat  and  give 
the  cleanest  possible  separation  of  matte  and  slag. 

The  following  is  a  good  example  of  modern  American 
furnaces  (Matthewson) : 

Grate:  6x16  ft.  Hearth:  20x100  ft.  (2000  sq.ft.). 
Flues:  3}  feet  wide.  Sand  bottom:  3  ft.  thick,  burnt  in 
one  layer.  Fuel  required:  57  tons  per  24  hours,  of  which 
6J  tons  are  recovered  by  treating  the  ashes  in  jigs.  Ore 
treated  averages  275  tons  per  24  hours  with  a  maximum  of 
330  tons.  It  is  taken  red  hot  from  the  roasters  every  80 
minutes  in  1 5-ton  lots  and  dumped  through  four  hoppers 


COPPER 


85 


upon  the  first  20  feet  of  the  hearth  back  of  the  fire  bridge. 
The  molten  matte,  8  in.  deep  on  the  hearth,  corresponds 
to  60-80  tons.  When  matte  is  needed  in  the  converter 
department,  it  is  drawn  into  ladles,  up  to  10  tons  at  a 
time.  Every  3  or  4  hours  30-40  tons  of  slag  are  withdrawn, 


FIG.  1 08. — Mansfeld  Furnaces. 

requiring  15  minutes  for  the  operation.     The  slag  is  granu- 
lated and  sluiced  to  the  dump. 

The  process  requires  little  hand  labor,  as  the  charge, 
dumped  in  the  hottest  zone  and  floating  upon  the 
matte,  distributes  itself,  and  then  flows  80  ft.  before 
reaching  the  taphole.  After  the  large  mass  has  been 
melted  down,  the  roof  and  side  walls  absorb  so  much 


86 


METALLURGY 


heat  that  at  the  end  of  80  minutes  a  new  charge  can  also 
enter  the  now  highly  superheated  chamber.  The  process 
thus  becomes  continuous  as  with  the  blast  furnace. 


FIG.  109. — American  Reverberatory  Furnace  (Peters).     Scale,  i  :  150. 

Since  the  reverberatory  furnaces  for  the  different  proc- 
esses in  copper  smelting  (matte  smelting,  black  copper 
smelting  and  copper  refining)  are  of  similar  design,  although 


COPPER 


87 


of  different  dimensions,  they  are  shown  together  on  pages 
85-87  (Figs.  108-110). 

1.  Oxidizing  Roasting  and 

2.  Smelting  for  Matte  are  often  carried  on  in  one  operation. 
This  is  exemplified  in  the  old  process,  kernel  roasting,  and  in 
pyritic  smelting,  a  process  brought  to  high  efficiency  during 
the  past  ten  years  in  Australia  and  in  the  United  States. 

Kernel  Roasting:  This  is  a  prolonged  heap  roasting,  the  pur- 
pose being  to  concentrate  the  copper  at  the  centre  of  the  slowly 
roasted  lump  of  ore.  By  this  method  of  roasting  there  takes 


FIG.  no. — Mathewson  Furnace,  Washoe  Smelter,  Anaconda.     Scale,  1:500. 


place  at  the  contact  of  the  oxidized  surface  and  the  unchanged 
sulphide  a  reaction  similar  to  that  which  occurs  in  matte 
smelting.  The  copper  is  taken  up  by  the  unaltered  sul- 
phide, which  results  in  a  constant  enrichment  of  the  centre 
in  copper,  leaving  a  crust  composed  of  iron  oxide.  At  the 
temperature  maintained  in  this  process,  part  of  the  copper 
is  unavoidably  converted  into  CuSC>4,  and  remains  as  such 
in  the  crust.  The  kernel  can  be  enriched  to  about  five  times 
the  copper  content  of  the  ore;  the  crust  usually  contains  about 
3%  Cu  as  CuSO4  and  i%  as  CuO. 

Pyritic  Smelting  (Metallurgie,  3,  1906).      True  pyritic  smelt- 
ing consists  of  an   oxidizing  smelt   without  using  any  fuel 


88  METALLURGY 

except  that  contained  in  the  ore  itself  (Fe  and  S).  It  is  used 
for  pyrite  carrying  copper  and  precious  metals,  but  with 
not  too  much  lead  (<io%)  or  zinc  (<io%  of  the  slag). 
For  ores  low  in  gold  there  should  be  0.3  to  0.5%  Cu;  for  ores 
rich  in  gold  (13  to  30  oz.  per  ton)  there  should  be  i  to  3% 
Cu  in  order  to  prevent  a  high  loss  in  gold. 

The  S  and  especially  Fe  are  the  sources  of  heat.  One 
part  by  weight  of  O  in  uniting  with  Fe  (to  FeO)  evolves  four 
times  the  amount  of  heat  as  when  uniting  with  S  (to  802). 
To  this  is  added  the  heat  of  neutralization  of  FeO  and  SiO2. 

Pyrite  (FeS2),  as  it  descends  through  the  furnace,  loses  S, 
but  often  has  not  attained  the  composition  FeS  at  the  beginning 
of  fusion.  On  entering  the  oxidation  zone,  however,  it  has 
usually  become  lower  in  S  than  FeS,  corresponding  to  about 
4FeS:Fe  or  3FeS:Fe.  The  presence  of  free  iron  in  this 
product  can  be  detected  with  a  microscope. 

The  amount  of  blast  must  be  sufficient  to  supply  the  SiO2 
with  as  much  FeO  as  possible  in  the  oxidizing  zone,  but 
not  enough  to  convert  the  Fe  to  Fe2Os.  Oxygen  oxidizes 
the  maximum  quantity  of  Fe  when  FeO  is  the  only  oxide 
formed.  If  the  ratio  required  to  form  the  singulo-silicate 
(FeO)2SiO2  is  reached,  the  full  heat  of  the  reaction  2FeO  + 
SiO2  is  available. 

If  there  is  a  lack  of  SiO2,  part  of  the  iron  will  be  converted 
into  the  higher  oxides  (Fe3O4  or  Fe2O3).  These  oxides  do 
not  themselves  form  silicates,  but  absorb  heat  on  being  dis- 
solved in  the  slag;  the  slag  becomes  cooler  and  more 
viscous. 

Hot  blast,  contrary  to  many  recommendations,  should 
be  avoided  in  most  cases.  It  favors  the  formation  of  acid 
slag  and  then  the  energy  of  saturation  of  the  SiO2  is  not  fully 
utilized.  Hot  blast  is  of  advantage  only  for  ores  which 
are  low  in  pyritic  material  and  require  considerable  addi- 
tional fuel. 

The  concentration  ratio  (ore:  matte)  becomes  greatest 
when  the  slag  most  nearly  approaches  a  singulo-silicate. 


COPPER  89 

Ores  of  such  composition  that  pure  pyritic  smelting  is  possible 
are  rare,  but  there  are  many  plants  that  have  reduced  the  fuel 
consumption  to  1-5%  of  the  total  charge,  whereas  it  was 
15-20%  when  using  roasted  ore. 

3.  Oxidizing  Roast  of  the  Matte  and 

4.  Matte  Concentration  by  Smelting  are  merely  repetitions  of 
operations  i  and  2,  and  the  object  and  apparatus  are  the  same. 
As  in  the  treatment  of  ores,  it  has  been  found  possible  to 
unite  processes  3  and  4  successfully  by  an  oxidizing  smelt. 
For  this  purpose  either  reverberatory  furnaces  or  converters 
are  used. 

Reverberatory  Smelting. 

This  has  been  already  sufficiently  described  under  2. 
Converting  Copper  Matte  is  not  only  carried  to  the  point  where 
the  matte  is  concentrated  to  Cu2S,  but  some  metallic    Cu 
may  be  formed.    It  is  merely  a  combination  of  the  roasting 
and  smelting  processes  of  matte  concentration,  carried  out 
by   blowing   compressed    air    through    the    molten  matte. 
The    matte    is    thereby    desulphurized  nearly  to  the   com- 
position  Cu2S,   the   iron   is   oxidized   to   FeO  and  slagged 
by  the  silicious  lining  of  the  converter.    After  the  matte  has 
been  converted  to  Cu2S  the  process  is  carried  out  in  different 
ways  and  this,  together  with  the  apparatus,  will  be  described 
under  "  (B)  Extraction  of  copper  "  (page  95). 
The  concentration  of  the  matte  aimed  at  in  these  enrichment 
processes  lies  between  72  to  78%   Cu.     Above    this    point   the 
reactions  soon  to  be  described  under  copper  extraction  take  place. 
This  is  due  to  the  fact  that  when  the  quantity  of  FeS  becomes 
small,    copper   separates   out    (copper   bottoms),    together   with 
silver  if  it  is  present  in  large  amount. 

A  third  roasting  and  smelting  is  to-day  seldom  performed, 
although  it  was  earlier  the  general  practice,  especially  in  Wales. 

Regarding  the  nature  of  copper  matte,  or  rather  of  the  different 
states  of  concentration,  there  was  until  recently  considerable  obscur- 
ity. Recently  several  researches  have  been  made  on  the  subject, 
of  which,  however,  only  one,  that  of  P.  Rontgen,  Aachen,  can  lay 


90  METALLURGY 

claim  to  great  accuracy.  According  to  this  work,  there  exist 
several  compounds  (at  least  three)  between  Cu2S  and  FeS,  one 
with  about  58%  Cu  corresponding  to  (Cu2S)3.(FeS)2,  one  with 
about  50%,  corresponding  to  Cu2S.FeS,  and  one  with  about 
33%  Cu,  corresponding  to  (0128)2- (FeS)  5.  Furthermore  there 
is  evidence  of  a  eutectic  between  Cu2S.FeS  and  (Cu2S)2.(FeS)5, 
corresponding  approximately  to  the  composition  Cu2S.2FeS: 

Cu2S.FeS  +  (Cu2S)  2.  (FeS)  5  =  3[Cu2S  +  2FeS]. 


90     ICK)^  Fe  8 
100    90     80     70     60     50     40     30     30     10     b$  Cu_S 

FlG.  III. 

It  is  interesting  to  note  that  if  we  imagine  the  first  compound 

(Cu2S)a.(FeS)2  to  be  more  highly  sulphurized  still,  we  come  to 

bournite  ore  (Cu2S)3.FeS.FeS2,  while  the  just-mentioned  eutectic 

corresponds  to  chalcopyrite  Cu2S.FeS.FeS2.    (See  Fig.  in.) 

Concentration  of  Copper  in  Aqueous  Solution  by  Leaching, 

can  be  substituted  advantageously  for  the  foregoing  smelting 

processes,  when  the  roasted  products  contain  copper  compounds 

(sulphate,  carbonate,  and  oxide)   that  are  readily  soluble  in 

water  and  acids,  but  not  when  the  relatively  insoluble  copper 


COPPER  91 

sulphides  are  united  with  other  sulphides.  Pure  copper  sul- 
phides may  be  converted  into  soluble  compounds  by  treatment 
with  different  agents  (see  below,  especially  ferric  and  cupric 
salts).  Ores  and  smelter  products,  on  the  other  hand,  in  which 
the  CuO  and  Cu2O  are  combined  with  other  oxides  (Fe2O3, 
SiC>2,  etc.)  or  in  which  the  Cu2S  is  combined  with  other  sulphides 
(FeS,  FeS2,  etc.,  in  chalcopyrite  and  bournite)  are  only  slightly 
attacked  by  the  customary  solvents.  Double  sulphides  may  be 
converted  by  roasting  at  a  low  temperature  (450-500°  C.)  into 
insoluble  Fe2Os,  soluble  Cu2O,  CuO,  CuSO4  and  unoxidized, 
but  free,  Cu2S.  Moreover,  according  to  a  recent  investigation 
by  O.  Frolich,  the  double  sulphide  may  be  decomposed  by 
heating  to  about  200°  into  the  simple  free  sulphides  without 
forming  any  oxides: 

Cu2S.FeS  =  Cu2S  +  FeS 

(cf.  Metallurgie,  1908,  5,  206). 

The  following  solvents  are  used : 

Water  for  Sulphates.  Since  in  weathering  and  roasting  some 
basic  sulphates  are  formed,  the  water  is  usually  acidulated 
if  it  is  not  directly  obtainable  as  acid  mine  water. 

Hydrochloric  Acid  is  used  ordinarily  in  connection  with  other 
salts,  e.g.  (NaCl),  to  increase  the  solubility  of  cuprous  and 
silver  chlorides.  At  the  Stadtberge  works  in  Lower  Marsberg, 
150  tons  of  acid  schist,  carrying  2%  Cu  as  carbonate  and  as 
copper  glance,  were  formerly  leached  daily  as  follows:  The 
ore,  crushed  to  6-15  mm.,  was  placed  in  shallow  trenches 
which  had  been  made  water-tight  with  clay  and  lined  with 
boards.  Here  it  was  treated  for  8-10  days  with  a  solution 
of  acid  and  NaCl,  whereby  copper  amounting  to  0.5%  of 
the  weight  of  the  ore  was  dissolved.  The  residue  was  then 
allowed  to  stand  in  the  air  for  from  8-10  weeks  in  well  venti- 
lated heaps  and  again  leached  with  salt  solution  (liquors 
containing  FeCl2  waste  from  the  precipitation  vats).  This 
dissolved  i%  Cu,  based  on  the  weight  of  the  ore.  After  the 
second  leaching  the  ore  was  transferred  to  the  dumps,  where 


92  METALLURGY 

it  was  again  subjected  to  the  action  of  the  atmosphere. 
Here  almost  all  of  the  remaining  copper  was  leached  out  by 
the  action  of  the  rain,  the  solution  being  removed  in 
drainage  trenches  and  conducted  to  the  precipitation  appar- 
atus. 

The  reactions  in  the  three  stages  of  the  leaching  are  as 
follows:  The  acid  dissolves  the  free  carbonate.  The  ore 
soaked  with  free  HC1,  FeCl2  and  FeCl3,  as  it  remains  in 
the  heaps  and  later  on  the  dumps,  undergoes  the  following 
changes : 

(1)  Cu2S  +  2FeCl3  =  2CuCl  +  S  +  2FeCl2. 

(2)  2FeCl2  +  2HC1  +  O  =  H2O  +  2FeCl3. 

The  Feds  thus  formed  reacts  again  as  under  (i).  The 
CuCl  may  also  be  converted  to  the  higher  chloride : 

(3)  2CuCl  +  2HC1  +  O  =  H20  +  2CuCl2. 
The  latter  now  works  in  a  similar  way  to  Feds. 

(4)  Cu2S  +  2CuCl2  =  S  +4CuCl. 

Sulphuric  Acid.  The  most  important  cases  where  sulphuric 
acid  is  utilized  have  already  been  described  under  Silver. 

Ferric  and  Cupric  Salts.  The  action  of  ferric  and  cupric  chlor- 
ides has  already  been  illustrated  under  hydrochloric  acid. 
Ferric  sulphate  acts  in  a  similar  way,  but  cupric  sulphate 
does  not,  because  cuprous  sulphate  is  so  unstable  that  it  is 
not  formed  under  the  conditions  that  prevail  in  leaching. 

In  the  Rio  Tinto  district  (Fig.  112),  where  pyrite  occurs 
with  free  Cu2S,  the  ore  is  arranged  in  long,  terraced  heaps  28 
ft.  high,  placed  side  by  side  and  provided  with  air  passages. 
The  heaps  are  sprayed  from  time  to  time  with  acid  mine 
water,  to  leach  out  the  CuSC>4  and  to  assist  in  the  sulphatizing 
action.  The  ferrous  sulphate  and  sulphuric  acid  contained 


COPPER 


93 


in  the  mine  water  and  in  the  heaps  apparently   come   from 
a  part  of  the  pyrite: 

(i) 


If  a  considerable  quantity  is  present,  it  may  be  oxidized 
to  ferric  sulphate: 


(2) 


FIG.  112. — Rio  Tinto  Leaching  Plant. 

The  Fe2(SO4)3  then  reacts  with  the  Cu2S  to  some  extent 
as  follows: 

(3)  2Fe2(SO4)3  +  Cu2S  =  2CuSO4  +  S +4FeSO4, 

A 

or,  together  with  the  atmospheric  oxygen  and  moisture, 

(4)  2Fe2 (SO4)3+Cu2S+H2O+3O  =  2CuSO4  +  4FeSO4+H2SO4. 
In  any  case  so  much  FeSO4  is    formed,  as  compared 


94  METALLURGY 

to  the  amount  of  H2SO4  present,  that  under  a  very  active 
oxidation  much  basic  sulphate  results  and  a  considerable 
quantity  of  hydrated  ferric  *oxide  forms  on  the  surface 
of  the  ore.  The  removal  of  the  copper  takes  place  so 
rapidly  that  much  FeS2  remains  unaltered  and  the  residue 
from  leaching,  still  containing  49  to  50%  S,  may  be  used 
in  sulphuric  acid  manufacture. 

Ferric    chloride  leaching  was  at   one   time   introduced 
into  the   Rio   Tinto   district,   but   the  practice   has   been 
abandoned  (Db'tsch  Process). 
Ferrous  Salts,  especially  FeCl2,  have  been  recommended  by 

Hunt  and  Douglas  in  leaching  ores  which  contain  copper  as 

oxide  or  carbonate: 

3CuO  +  2FeCl2  =  Fe2O3  +  CuCl2  +  2CuCl. 

As  an  argument  in  favor  of  the  process  it  was  pointed  out  that 
CuCl  solution  requires  less  iron  for  precipitating  the  copper 
than  CuCl2  or  CuSO4  solutions. 

Ores  which  cannot  be  leached  directly  are  prepared  under  some 
circumstances  by  subjecting  them  to  a  chloridizing  or  a  sulphat- 
izing  roast.     In  both  cases  an  oxidizing  roast  is  carried  out;  in 
the  first  case  chlorides  (MgCl2,  NaCl,  etc.)  are  added  and  chlorine 
is  made  available  by  the  aid  of  SO2  and  O,  and  in  the  second 
case  the  temperature  is  kept  as  low  as  possible   (450  to  500°  C). 
A  PURIFICATION  OF  THE  LEACHED  LIQUOR  is  necessary  if 
the  plant  also  sells  copper  as  blue  vitriol.     The  chief  impur- 
ity is  FeSO4,    together   with  small    quantities    of  Sb,  As, 
etc.     All  these  are  substances  easily  precipitated  if  copper 
oxide  (roasted  matte)  is  introduced  while  air  is  being  blown 
through  the  hot  solution  for  the  purpose  of  oxidizing  FeSO4 
to  Fe2(SO4)3. 

2CuO  +  2FeSO4  +  O  -  2CuSO4  +  Fe2O3. 

This  process  was  brought  to  a  high  state  of  efficiency  by 
Ottokar  Hofmann  in  Kansas  City  (Metallurgie,  1907,  4,  582). 


COPPER  95 

(B)   Extraction  of  Copper  from  Intermediary 

Products 

In  the  concentration  processes  described  under  A,  the  copper 
remained  either  in  the  form  of  sulphide  low  in  iron  (Cu2S) 
often  carrying  precious  metals,  or  in  the  form  of  an  aqueous 
solution.  The  concentrated  matte  is  commonly  converted  into 
copper  by  smelting  processes,  among  which  the  so-called  Reaction 
Process  is  most  commonly  employed,  although  the  Roast-reduction 
Process  has  found  favor  if  a  matte  containing  silver  is  to  be 
desilverized  by  leaching  between  the  roasting  and  smelting 
operations.  If  the  copper  has  been  concentrated  in  the  wet 
way,  obviously  only  precipitation  methods  come  into  consideration. 

Reaction  Smelting  is  carried  out  either  in  reverberatories  or  in 
converters,  and  causes  the  separation  of  copper  according  to 
one  of  the  following  reactions : 

With  the  sulphides  of  base  metals,  the  oxidation  always 
takes  place  so  that  the  metal  oxide  and  SO2  are  formed: 

Cu2S  +  30  =  Cu2O  +  SO2. 

The  higher  oxide  is  not  formed  in  the  presence  of  reducing 
agents  (Cu2S),but  the  latter,  on  the  contrary,  reduces  the  Cu2O: 

2Cu2O  +  Cu2S  -  3Cu2  +  SO2. 

The  reverberatory  furnaces  used  for  this  process  are  not  so 
large  as  the  newer  smelting  furnaces,  but  like  the  latter  they  have 
a  sand  bottom  and  are  provided  with  tuyeres  beside  the  firebox 
for  blowing  air  upon  the  metal  bath  (cf.  page  85-87). 

In  1880  Manhes  in  France  and  one  of  his  assistants  at  the 
Parrot  Copper  Company  of  Butte,  Montana,  succeeded  in  blow- 
ing copper  matte  directly  to  metallic  copper  in  a  converter  similar 
to  that  used  in  the  Bessemer  process.  The  success  was  due  to  a 
change  in  the  method  of  admitting  the  air.  The  previous  failures 
had  been  due  to  using  a  converter  of  exactly  the  same  pattern  as 


96 


METALLURGY 


that  used  in  steel  manufacture  without  taking  into  account  the 
different  conditions  governing  the  working  up  of  the  raw  materials 
used  in  copper  smelting.  In  the  copper  Bessemer  process  a  matte 
carrying  40  to  55%  Cu  is  commonly  used.  In  comparison  with 
the  iron  Bessemer  process  (see  Iron)  the  following  differences  may 
be  noted. 


Copper  Matte. 

Iron. 

45  to  60%  oxidizable  substances  of 
low  calorific  power. 

3  to  6%  impurities  of 
power. 

high  calorific 

Formation  of  large  amounts  of  strong- 
ly basic  FeO  which  slags  with  the 
converter  lining. 

Quantity  of  slag  small. 

Converter  lining  soon  destroyed  (9 
charges). 

Converter    lining    lasts 
(200-250  charges). 

a    long  time 

Three  different  materials:  slag,  con-  j  Two     materials:    small    amount     of 
stantly  increasing;  matte,  constant- 


ly decreasing;   metal,  in  the  second 
stage  constantly  increasing. 


Large  amount  of  metal. 


Copper:    good     conductor    of    heal. 

Specific     heat     0.155      (liquid,      see 

Glaser,  Metallurgie,  1904,  I,  126). 


Iron:  poorer  conductor.  Specific 
heat  0.1665  (liquid,  see  Oberhoffer. 
Metallurgie,  1907,  4,  495). 


Only  when  Manhes  arranged  the  tuyeres  so  that  the  air  no 
longer  passed  through  the  separated  metallic  copper  was  the 
blowing  successful. 

MANHES'  CONVERTER.  This  consists  of  a  cylinder  on  a 
horizontal  axis;  the  tuyeres  are  arranged  in  a  row,  in  the 
cylinder  mantle,  being  placed  parallel  to  the  axis,  so  that 
by  rotating  the  drum  the  tuyeres  can  be  made  to  enter  the 
matte  at  different  depths.  This  cylindrical  type  has  been 
successful  at  both  the  Anaconda  and  Copper  Queen  (Bisbee) 
plants.  Fig.  113  shows  a  modern  arrangement  of  such  a 
plant. 

STALMANN  constructed  a  converter  of  rectangular  cross- 
section.  These  converters  have  been  introduced  by  Sticht  at 


COPPER 


97 


FIG.  113. — Horizontal  Converters  at  the  Anaconda  Plant. 


FIG.  114. — Converter  Department,  Mount  Lyell  M.  &  R.  Co.,  Tasmania. 


METALLURGY 

the  large  plant  of  the  Mount  Lyell  Mining  and  Railway  Co., 
Australia.     Fig.  114  shows  the  converter  department  of  these 


c/: 


works.      Figs.  115,  116  and  117  also  show  sketches   of   the 
converters  as  made  by  Mr.  Sticht. 

Upright  converters,  resembling  the   Bessemer   type,    have 
been  successfully  used    at    the  Parrot  and  Anaconda  plants. 


COPPER 


99 


The  tuyeres  are  placed  in  a  semicircle  above  the  bottom   on 

one  side  of  the  converter  (Figs.  118-121,  p.  100). 

The    converter    lining,    consisting    of    17-28%    clay  and 

83-72%  quartz  (gold  bearing),  is    tamped    into  a  thickness 

of  about  1 8  in. 

The  blowing  takes  place  in  two  stages:   (i)  Oxidation  of 

the    FeS    and    slagging    of    FeO,    usually    requiring    30-50 

minutes,  and  (2)  blowing  to  separate  the  copper,  requiring 

30-50  minutes  more. 
The  Roast-Reduction  Process : 

This  consists  of  the  following  operations: 
i.  OXIDIZING  ROAST  (dead  roast)  for  the  purpose  of  removing 


FIG.  117. 
Stalmann  Type  of  Converter  at  Mount  Lyell. 


Designed  by  R.  Sticht. 


as  much  of  the  S  as  possible  and  converting  the  Cu  into  CuO. 
(Apparatus:  see  A,  p.  61  to  71.) 

2.  REDUCING  SMELT  in  a  shaft  furnace  or  a  reverberatory,  with 
other  rich  copper-bearing  materials  if  desired.  This  is  very 
successfully  used  at  the  Mansfeld  works  in  conjunction  with 
the  Ziervogel  process.  Reverberatory  furnaces  are  used  at 
this  plant. 

Precipitation  of  Copper  from  solutions  of  the  chloride  and  sul- 
phate is  almost  always  accomplished  by  iron.  In  the  Stadt- 
berge  Works  (page  91)  wooden  vats  are  used  in  which  the 
iron  in  the  form  of  thin  turnings  rests  upon  a  wooden  frame 


100 


METALLURGY 


and  the  solution  is  agitated  by  wooden  paddles.  At  Rio 
Tinto  the  copper  is  precipitated  in  trenches  in  which  cast 
iron  bars  have  been  placed.  (Fig.  122). 


FIG.  11 8. 


FiG.119. 


FIG.  120. 

Converters  at  the  Anaconda  Plant. 


FIG.  121. 


The    precipitated    copper   is   treated    in    jigs   to   remove 
large    pieces    of   iron    and    in    washing    drums    to   remove 


COPPEk 


101 


fine  particles  of  iron  and  iron  oxide.     It  is  then  filtered  and 
dried. 

If  the  solution  carries  silver,  part  of  the  copper  (one-third 
to  three-eighths)  is  first  precipitated,  carrying  with  it  the 
silver,  then  the  remaining  copper  is  precipitated,  silver 


m 


FIG.  122. — Rio  Tinto  Precipitating  Trenches. 

free,  in  a  second  vat.     The  two  lots  are  treated  differently 
(see  below). 

(C)   Copper  Refining 

Pure  copper,  or  refined  copper,  is  made  to-day  almost  entirely 
from  crude  copper.  The  attempts  to  produce  pure  copper 
directly  from  the  ore  have  thus  far  proved  a  failure,  and  the 
same  is  true  of  most  attempts  to  produce  pure  copper  from 
matte.  It  is  only  recently  that  the  difficulties  hitherto  met  with 
in  the  latter  process  have  been  overcome. 


102 

Crude  copper  may  contain  the  following  impurities:  S,  As,  Sb, 
Zn,  Pb,  Fe,  Ni,  Co,  Ag,  Au.  If  precious  metals  are  not  present 
the  following  process  may  be  used : 

Furnace  Refining  consists  of  an  oxidizing  followed  by  a  reduc- 
ing fusion. 

1.  OXIDIZING  FUSION.     A  reverberatory  furnace  is  usually 
used  for  this  purpose,  the  object  being  to  eliminate  the 
oxidizable  impurities    from  the  copper.     The  S,  As,  Sb, 
Zn,  and  Pb  are  partly  volatilized  as  oxides  and  partly 
slagged;  any  Fe  and  Ni  are  also  slagged.     The  S  is  present 
in  impure  copper  largely  as  Cu2S,  and  is  removed  only 
after  the  greater  part  of  the  other    impurities    has    been 
eliminated.     It  is  oxidized  principally  through  the  agency 
of  Cu2O  after  a  certain  quantity  of  the  latter  has  been 
dissolved   in    the   copper.     When    this   point   is   reached, 
the  evolution  of  SC>2  takes  place  rapidly  (boiling);    when 
the  reaction  has  nearly  ceased,  a  wooden  pole  is  plunged 
into  the  metal. 

The  gases  escaping  from  the  submerged  end  of  the  pole 
agitate  the  metal,  thus  allowing  the  SO2  to  escape  more 
easily  and  at  the  same  time  the  spurting  copper  is  further 
oxidized  (dense  poling).  The  copper  resulting  from  this 
treatment  contains  much  Cu2O  (set  copper).  The  result- 
ing slag  runs  high  in  copper  and  goes  back  to  the  smelting 
furnace.  Set  copper  is  very  brittle  and  unfit  for  the 
ordinary  purposes  for  which  copper  is  utilized.  It  is 
further  refined  as  follows: 

2.  REDUCING  FUSION.     The  molten  copper  is  heated  with 
a    reducing    flame    and    kept    covered    with    charcoal; 
a  wooden  pole  (green  wood)  is  introduced  and  poling  con- 
tinued until  the  dissolved  Cu2O  has  been  reduced  by  the 
hydrocarbons  in  the  pole  and  the  charcoal  in  the  covering. 
This  process  is  complete  when  a  thin  test-bar  can  be  bent 
through    1 80°   without   showing  a   crack.     When   broken 
by  repeated  bending,  the  fracture  must  show  a  fibrous, 
salmon-red  texture  with  silky  luster. 


COPPER 


103 


The    refined  copper   is  cast  from  wrought-iron  ladles, 
which  are  coated  with  chalk  or  clay,  into  cast  iron  or  copper 


CASTING  MACHINE, 
WALKER  SYSTEM. 


FIG.  124. 

molds.  In  large  plants  it  is  run  directly  from  the  furnace 
through  troughs  into  mechanically  dumping  molds  (Figs. 
123-124). 


104 


METALLURGY 


Electrolytic    Refining    is  almost   universally   used   for  copper 
containing  precious  metals.     It  is  carried  out  in 
ELECTROLYZING  VATS   made  of  wood  but  lined  with  lead. 


FIG.  125. 


Anode 


Cathode 


I 


FIG.  126. 


FIG.  127. 


THE  ANODES  are  cast  plates  of  unrefined  copper  (Figs.  125, 
126,  and  127)  that  contains  precious  metals. 

THE  CATHODES  are  thin  sheets  of  copper  made  by  deposit- 
ing the  metal  electrically  on  lead  or  greased  copper  plates 
from  which  it  can  be  easily  removed. 


COPPER  105 

THE  ELECTROLYTE  is  an  acid  solution  of  copper  sulphate 
containing  12-15%  CuSO4.5H2O  and  5-10%  H2SO4. 

DURING  THE  PROCESS,  it  is  important  that  the  constantly 
decreasing  quantity  of  free  acid  and  the  constantly  increas- 
ing quantity  of  Cu  in  the  electrolyte  does  not  get  outside 
the  above  limits.  More  copper  is  dissolved  from  the  anode 
than  corresponds  to  the  current  density,  because  of  the 
action  of  the  atmospheric  oxygen  : 

Cu  +  O  +  H2SO4  =  CuSO4  +  H2O. 

The  above  reaction  takes  place  under  normal  conditions 
only  at  the  surface  of  the  anode.  If,  however,  the  proc- 
ess is  carried  on  at  a  low  current  density,  a  part  of  the 
copper  as  it  goes  into  solution  from  the  anode  may  become 
incompletely  charged,  assuming  a  univalent  instead  of  a 
bivalent  charge;  in  other  words,  the  solution  then  con- 
tains cuprous  ions,  but  these  are  precipitated  again  at 
the  anode  : 


It  follows,  therefore,  that  the  insoluble  particles  remaining 
at  the  surface  of  the  anode,  which  fall  to  the  bottom  as 
slime,  are  likely  to  contain  considerable  copper.  The 
latter,  however,  will  be  easily  dissolved  if  the  electrolyte  is 
agitated  with  air  so  that  the  concentration  at  the  electrodes 
is  equalized. 

THE  TEMPERATURE  may  be  raised  to  40°  C.  (104°  F.);  there 
is  no  advantage  in  raising  the  temperature  above  this 
point,  if  the  quality  of  the  cathode  copper  is  taken  into 
consideration. 

THE  CURRENT  DENSITY  may  vary,  according  to  the  quality 
of  the  anode  copper,  between  4  and  15  amperes  per  sq.ft. 

THE  POTENTIAL  should  be  0.1-0.3  volt. 

THE  REACTIONS  DURING  ELECTROLYSIS  are  as  follows:  The 
anode  copper  carrying  a  positive  charge  goes  easily  into 
solution  and  passes  to  the  cathode,  where  it  loses  its 


106  METALLURGY 

charge  and  is  deposited.  Of  the  impurities  present  in 
anode  copper,  As,  Sb,  Bi  and  Sn  go  partly  into  solu- 
tion, but  are  again  partly  precipitated;  Ag,  Au,  Pt,  and 
Cu2S  are  insoluble;  Fe,  Zn,  Ni  and  Co  go  into  solution, 
but  are  not  deposited  on  the  cathode.  The  insoluble 
substance  falls  in  a  finely-divided  form  to  the  bottom 
near  the  anode,  while  the  soluble  substances  that  are  not 
deposited  remain  and  gradually  contaminate  the  electrolyte, 
so  that  it  must  be  drawn  off  and  its  valuable  constituents 
recovered. 

PRODUCTS  OF  ELECTROLYTIC  REFINING: 
Cathode  copper  (electrolytic  copper) ,  97-99%  of  the  anode 
copper; 

Precious  metals,  ] 

Bismuth,  tellurium,  etc.,  [•    from  the  anode  mud. 

Blue  vitriol. 

Impure     cement     copper  ] 

/         .    IN      r     n       1        from  the  anode  mud  and  the 
(arsenical) ,    finally  also  \ . 
......  impure  electrolyte. 

nickel  vitriol. 

Electrolytic  Treatment  of  Copper  Matte  (Process  of  Borchers, 
Franke  and  Giinther)  has  been  carried  out  since  1907  in 
an  experimental  plant  at  the  Mansfeld  works.  In  previous 
unsuccessful  attempts  low-grade  matte  was  used,  but  in 
this  case  matte  carrying  72%  Cu  and  upward  is  employed. 
In  its  essential  features,  the  method  of  working  is  the  same 
as  that  in  the  electrolytic  refining  of  metallic  copper  that 
carries  precious  metals.  The  principal  differences  in  pro- 
cedure and  the  products  are  as  follows: 
ANODES,  cast  plates  of  concentrated  matte  (Cu2S). 
POTENTIAL:  higher  than  in  treating  metallic  copper,  viz., 

0.75  volt. 

REACTIONS  DURING  ELECTROLYSIS:  with  the  proper  cur- 
rent density,  the  copper  is  dissolved  as  when  the  anode 
is  crude  copper,  and  the  sulphur  from  the  Cu2S  remains 
behind  in  the  free  state. 


COPPER  107 

The  other  constituents  of  the  matte  act  the  same  as  in 
metal  refining.  If  the  current  density  is  too  low,  only  half 
of  the  copper  is  dissolved,  the  balance  remaining  as  blue 
CuS. 

PRODUCTS  OF  THE  PROCESS:     Besides  the  products  obtained 
from  treating  metal,  there  is  in  addition, 
Sulphur,  which  is  recovered  as  yellow  crystals  on  drying 

the  anode  mud. 
Although  this  process  requires  more  power  it  possess  the 

following   advantages : 
ELIMINATION    OF    SMELTING    the   concentrated    matte   to 


FIG.  128.  FIG.  129. 

black  copper,  and  elimination  of  the  damage  due  to  the 
fumes  from  such  work; 

LESSENING  OF  LOSSES  in  precious  metals,  which  are 
greater  in  smelting  for  black  copper  than  in  any  other 
stage  of  copper  treatment. 

RECOVERY   OF    THE    SULPHUR    contained  in  the    matte. 
Properties  of  Refined  Copper : 
Specific  gravity:  8.94. 

COLOR:  yellowish-red,  brilliant  luster.  x 

MECHANICAL  PROPERTIES:    of  medium  hardness  and  tensile 
strength,  very  ductile. 


108  METALLURGY 

STRUCTURE    of    cast  and  electrolytic  copper,  granular;    of 
rolled  and  hammered  copper,  fibrous  (see  Figs.  128-129), 
MELTING-POINT:  1084°  C.  (1983°  F.). 
BOILING-POINT:    is  said  to  be  in  the  vicinity  of  2100°  C. 

(3815°  F.). 
THERMAL  AND  ELECTRICAL  CONDUCTIVITY:  nearly  as  great 

as  Ag  (0.96). 

ALLOYS  easily  with  Mg,  Al,  Mn,  Zn,  Cd,  Co,  Ni,  Hg  and  the 
precious  metals,  slightly  with  Fe,  Mo,  W,  Cr  if  the  metal 
is  pure,  more  readily  in  the  presence  of  Si.  Copper  also 
alloys  easily  with  its  own  compounds,  especially  with 
Cu2O  and  Cu2S,  and,  when  in  the  molten  state,  it  alloys 
somewhat  with  H,  CO  and  SO2. 

CHEMICAL  BEHAVIOR.  In  the  solid  state  it  is  fairly  resistant 
to  dry  oxygen  at  the  lower  temperatures,  but  above  400°  C 
it  is  easily  oxidized.  Crusts  of  CuO  and  Cu2O  (copper 
hammer  scale)  form  on  the  surface.  In  a  damp  atmosphere 
it  easily  forms  basic  salts  with  O  and  weak  acids  (verdigris) . 
In  the  molten  state  copper  has  a  great  affinity  for  S,  the 
the  highest  of  any  metal  except  manganese.  Of  the 
acids,  only  the  oxidizing  acids,  HNO3,  hot  concentrated 
H2SO4,  and  aqua  regia,  serve  as  solvents.  With  access  of 
air,  dilute  and  weak  acids  attack  copper.  Copper  does 
not  evolve  hydrogen  from  acids,  as  its  solution  tension 
toward  hydrogen  is  —0.329  volt. 


BISMUTH 

Sources 

Natural  Sources: 

FREE,  with  small  quantities  of  S  and  As. 
MINERALIZED  as 

OXIDE,  Bi2Os  in  bismuth  ochre,  and  as 
SULPHIDE,  Bi2S3,  in  bismuth  glance.    The  sulphide  occurs 
free,  in  solution  or  in  combination,  with  sulphides   or 
arsenides  of  cobalt,  nickel,  copper,  lead  and  silver. 
Other  Sources : 

LITHARGE  and  other  products  from  the  cupelling  process. 
REFINERY  SLAG  from  silver  refining.     Residues  from  the 
preparation  of  pure  bismuth  salts.     These  may  be  pure 
products  for  which  there  is  no  demand  or  salts  that  have 
become  contaminated  with  impurities. 

(A)   Concentration  Processes 

Mechanical  Concentration  consists  usually  in  hand  picking. 

Chemical  Concentration  is  used  in  working  up  litharge  that 
contains  bismuth.  In  the  section  on  Silver,  it  was  stated 
that  at  the  beginning  of  the  cupellation  process,  when  bis- 
muth is  present  in  the  bullion,  only  the  lead  is  oxidized; 
later,  as  the  bismuth  becomes  more  concentrated  in  the  metal 
bath,  it  is  also  oxidized,  forming  Bi2Os,  which  enters  the 
litharge.  The  latter  seldom  reaches  a  concentration  of  more 
than  i  to  2%  Bi.  It  is  prepared  for  the  regular  Pb-Bi  sep- 
aration by: 

i.  Reducing  Fusion  to  form  a  Bi-Pb  alloy,  which  is  low  in  Bi 
at  the  start. 

100 


110  METALLURGY 

2.  Oxidizing    Fusion,  whereby  litharge  is  formed  which  is  at 
first  low  in  bismuth  but  later  becomes  richer.     The  different 
lots  of  litharge  are  kept  separate  and  operations  i  and  2  are 
repeated   with  the   litharge  low  in   Bi.     The  rich  litharge 
(more  than  20%  Bi)  is  treated  by  leaching  and  precipita- 
tion, usually  by 

3.  Agitation  with  Dilute  HC1  (15%)  accompanied  by  blowing 
steam  through  the  solution;    this  results  in   the   changing 
of  the  oxide  to  chloride. 

APPARATUS:     earthenware    jars    with    wooden    covers    and 

spigots. 
PRODUCTS:     an    acid    Bids    solution    containing  a   small 

amount   of   PbCl2   and,  in  suspension,  some   undissolved 

PbCl2  which  is  removed  by 

4.  Filtration. 

5.  Stirring  the    BiCl3  solution  into  water  and  neutralization 
of  most  of  the  free  acid  with   Ca(OH)2  CK  Na2CO3.     This 
causes  the  hydrolysis  of  the  Bids: 

BiCl3 + H20  =  BiOCl  +  2HC1. 

The  BiOCl  is  precipitated,  while  the  PbCl2,  not  previously 
removed  by  filtration,  remains  dissolved  in  the  weakly  acid 
solution.  The  BiOCl  is  filtered,  dried,  and  is  reduced  to 
Bi  or  is  used  in  fire  refining  bismuth  that  carries  lead. 

(B)   Extraction  of  Bismuth 

Liquation :  Ores  carrying  native  bismuth  were  formerly  subjected 
to  liquation,  in  crucibles  or  retorts,  to  remove  the  greater  part 
of  the  bismuth  without  melting  the  gangue.  The  process  can 
be  used  only  on  rich  ores  and  the  gangue  usually  retains  4-5% 
Bi.  The  gangue  is  re-treated  by  one  of  the  following  processes : 

Reduction  Process:  applicable  to  oxidized  ores,  intermediate 
products  and  residues.  Among  the  two  latter  are  bismuth 
salts,  e.g.,  the  oxychloride.  Because  of  the  volatility  of  Bi2O3 


BISMUTH  HI 

and  of  Bi  itself,  and  because  of  the  ready  reducibility  of  Bi2O3, 
the  temperature  should  be  kept  low  during  this  process.  It 
is  necessary,  therefore,  to  produce  a  slag  of  the  lowest  possible 
melting  point  in  order  that  it  may  be  as  free  from  bismuth  as 
possible.  The  silica  content  of  such  a  slag  should  lie  between 
a  singulo  and  bi-silicate,  in  fact  as  near  to  the  latter  as  possible. 
The  bi-silicate  slags  used  in  other  smelting  processes  (Cu,  Pb), 
which  contain  chiefly  FeO  and  CaO  as  bases,  are  too  infu- 
sible for  this  process.  In  this  case,  therefore,  Na2COs  is  used 
as  a  flux,  besides  materials  carrying  FeO  and  CaO.  The  high 
price  of  Bi  permits  the  use  of  this  somewhat  expensive  flux, 
especially  as  the  fluidity  of  the  slag  is  such  that  the  Bi2O3  and 
the  reducing  carbon  quickly  sink  through  it,  thus  reducing  to  a 
minimum  the  loss  by  dusting  and  volatilization.  If  basic 
bismuth  salts,  such  as  BiOCl,  are  to  be  treated,  they  should 
not  be  added  directly  to  the  melted  charge  because  of  their 
volatility.  They  should  rather  be  stirred  into  wet  Ca(OH)2 
or  Na2COs,  dried  and  calcined  for  a  time  in  order  that  the  Cl 
may  be  completely  taken  up  by  the  Ca  or  Na. 

APPARATUS:     the   process   may   be   carried    out    either    in 

crucibles  or  in  reverberatory  furnaces. 

Melting  in  Crucibles  does  not  require  highly  refractory  materials, 
and  hence  well-baked  clay  crucibles  may  be  used,  although 
it  should  be  borne  in  mind  that  the  slag  produced  in  the 
process  will  dissolve  basic  as  well  as  acid  oxides.  For  mak- 
ing these  crucibles,  which  may  be  done  as  part  of  the  process, 
it  is  important  to  use  as  a  foundation  a  clay  which  has  been 
well  baked  and  subsequently  ground.  Such  material  is  mixed 
thoroughly  with  the  least  possible  amount  of  fresh  clay  for  a 
binder.  The  crucibles  are  usually  formed  by  hand.  For 
heating,  a  simple  air  furnace  may  be  used,  but  for  large 
plants  small  continuous  kilns  are  preferred,  since  they  allow 
a  better  control  of  the  temperature  and  are  not  so  hard  on 
the  crucibles.  The  latter  is  an  important  point  because  the 
use  of  a  graphite  crucible,  which  more  easily  resists  temper- 
ature changes,  is  prohibited.  The  corrosive  action  of  the 


112 


METALLURGY 


FIG.  133. 


Borchers*  Continuous  Kiln. 


BISMUTH  113 

slag  permits  the  use  of  a  crucible  only  once,  and  the  expense  of 
graphite  crucibles  would  greatly  increase  the  cost  of  the 
process.  The  common  clay  crucibles  fulfil  all  requirements 
if  they  are  carefully  handled.  This  is  difficult  in  air  fur- 
naces, in  which  the  fuel  conies  in  direct  contact  with  the 
crucible,  where  a  cold  piece  of  coke  touching  the  hot  wall 
•may  cause  it  to  crack.  A  small  continuous  kiln  designed 
by  Borchers,  with  six  chambers  heated  with  producer  gas, 
which  is  generated  in  a  producer  attached  to  the  kiln,  com- 
pletely overcomes  the  difficulties  encountered  in  the  air 
furnace  and  also  results  in  a  saving  of  fuel. 

In  the  continuous  kiln  shown  in  Figs.  130  to  133,  the  hot 
gases  are  conducted  from  the  producer  through  the  main 
flue  which  runs  along  the  center  of  the  upper  part  of  the  fur- 
nace. From  this  flue  the  gases  can  be  conducted  by  means 
of  fl-pipes  of  sheet  iron,  fitting  into  corresponding  faucet 
pipes,  to  a  branch  flue  for  each  chamber.  Sliding  doors 
have  not  been  used  here,  as  they  cannot  be  fitted  tightly 
enough  to  prevent  the  gas  from  entering  chambers  which 
are  cooling.  From  each  of  the  branch  canals  a  number  of 
small  slits  enter  the  corresponding  chamber;  between  the 
slits  are  openings  from  an  air  flue  lying  beneath.  The  com- 
bustion begins  in  the  chambers  above  the  crucibles  and  the 
hot  gases  surround  the  crucible  from  top  to  bottom.  The 
crucibles  themselves  rest  upon  supports,  between  which  are 
openings  that  allow  the  gases  to  pass  into  the  space  below 
and  thence  through  connecting  flues  to  the  next  chamber, 
where  they  warm  the  filled  crucibles.  The  gases  usually 
pass  through  a  second  chamber  before  leaving  the  furnace. 
The  air  is  not  admitted  directly  to  the  combustion  chamber 
but  is  first  passed  through  two  chambers  containing  crucibles 
filled  with  hot  finished  product.  In  cooling  these  crucibles 
the  air  becomes  pre-heated.  From  Figs.  130  to  133  it  is 
seen  that  six  chambers  are  so  arranged  that  an  uninterrupted 
continuous  process  can  be  carried  on.  As  soon  as  the  con- 
tents of  one  chamber  are  melted,  gas  and  air  are  admitted 


114 


METALLURGY 


•• 

f  - 


into  the  next  chamber,  which  has  already  been  strongly  pre- 
heated, and  the  products  of  combustion  are  conducted 
through  an  intermediate,  freshly-charged  chamber.  The 
chamber  in  which  the  air  first  entered  is  now  closed,  the 
crucibles  are  removed,  and  a  new  charge  is  put  in  place. 
Melting  in  Reverberatory  Furnaces.  In  constructing  the  rever- 
beratory,  it  should  be  borne  in  mind  that  bismuth  in  the 


FIG.  135. 


H 


FIG.  136. 

Borchers'  Reverberatory  Furnace. 


FIG.  137. 


molten  state  easily  leaks  out  through  the  finest  cracks  or  imper- 
fections in  the  walls.  Furnaces  of  earlier  construction,  having 
heavy  brickwork  supports  for  the  hearth,  are  being  discarded 
because  they  absorb  large  quantities  of  bismuth,  thus  dimin- 
ishing the  output  until  the  furnace  is  torn  down.  Even  then  the 
recovery  of  the  metal  from  the  hearth  material  is  tedious  and 
costly.  In  plants  designed  and  constructed  by  Borchers,  fur- 
naces have  been  built  with  movable  iron  hearths  lined  with 
half  a  course  of  bricks  (Figs.  134  to  137).  The  advantages 


BISMUTH  115 

of  these  hearths  are  a  better  and  quicker  extraction  of  the 
bismuth  and  easier  access  to  the  hearths  for  repairs  to  the 
lining,  without  tearing  down  the  side  walls.  The  hearths  can 
be  easily  changed.  Moreover,  firebox  and  flue  should  be 
isolated  from  the  hearth  walls  by  an  air  space  or  a  water- 
cooled  plate,  as  otherwise  the  bismuth  would  flow  into  the 
firebox  or  flue  through  the  two  ends  that  are  heated  by  the 
fire  or  flue  bridge. 

At  the  beginning  of  the  process  in  the  reverberatory  furnace, 
some  slag  is  first  melted  and  then  the  fresh  charge  is  ..intro- 
duced.    This  should  if  possible  sink  at  once  t<\$\  Bottom, 
to  prevent  loss  by  dusting  and  volatilization. 
Precipitation  of  Bismuth  is  based  upon   the  decomposition  of 
Bi2S3  by  means  of  Fe. 


As  the  equation  shows,  this  process  is  applicable  to  sulphide 
ores.  Apparatus  and  operations  are  the  same  as  in  the  reduction 
process.  If  the  ores  contain  substances  which  may  be  con- 
verted into  matte  (Cu,  Co  and  Ni)  or  into  speiss  (As,  Sb)  the 
slag  should  not  be  as  acid  as  in  the  reduction  process.  The 
silica  content  may  lie  nearer  that  of  a  singuto-silicate. 

(C)    Bismuth  Refining 

Crude  bismuth  usually  contains  As  and  Sb,  often  Pb,  and 

occasionally  precious  metals.     Of  these  impurities,  As  and  Pb, 

also  Sb  if  it  is  present,  in  small  quantities,  may  be  removed  by  an 

Oxidizing  Fusion.     The  choice  of  an  oxidizing  agent  depends 

upon  the  nature  and  amount  of  the  impurities.     If  much 

lead  is   present   this   must   first   be  removed,  as   otherwise 

lead  compounds   (plumbates)    are  formed  which  favor  the 

oxidation  of  large  quantities  of  bismuth.     The  lead  is  best 

removed  by 

MELTING  WITH  BiOCl  in  iron  kettles  under  a  neutral  cover 
of  NaCl  and  KC1,  or  by 


116  METALLURGY 

MELTING  WITH  NaOH  with  the  addition  of  NaNO3,  so 
that  As  can  be  oxidized  to  Na3AsC>4  in  the  same  appa- 
ratus. 

In  the  above  operation  the  NaCl  and  KC1  or  the  NaOH 
is  first  melted,  then  the  metal  added,  followed  by  the 
oxidizing  agent,  BiOCl  or  NaNO3,  which  is  stirred  in  with 
an  iron  spatula. 

When  the  refining  is  completed,  an  iron  hook  is  low- 
ered in  the  molten  metal  and  the  kettle  is  cooled.  After 
the  mass  has  solidified,  the  slag  is  dissolved  in  hot  water, 
th^V^ttle  is  warmed  to  free  the  metal  from  the  walls, 
andv  Ais  is  then  lifted  out  by  means  of  the  imbedded 
hook. 

Melting  with  Sulphur  and  Na2CO3  or  K2CO3  is  used  in  excep- 
tional cases  when  the  metal  is  high  in  Sb.     The  apparatus 

and  procedure  are  the  same  as  above. 
Electrolysis  is  used  when  the  bismuth  carries  precious  metals. 

VESSELS  :  stone  jars. 

ANODES:  plates  or  blocks  of  Bi  carrying  precious  metals. 

CATHODES:  pure  Bi. 

ELECTROLYTE:  HNO3  and  Bi(NO3)3,  or  HC1  and  BiCl3,  in 
aqueous  solution. 

CURRENT  DENSITY:   15-30  amperes  per  sq.  ft. 

POTENTIAL:  0.5-1  volt. 
Properties  of  Refined  Bismuth : 

SPECIFIC  GRAVITY:  9.74-9.8. 

COLOR:   bright  gray,  slightly  reddish  luster. 

MECHANICAL  PROPERTIES:  very  brittle,  easily  pulverized. 

STRUCTURE:  large- faced,  isometric,  crystalline  grains  having 
a  characteristic  dendritic  structure  (Fig.  138). 

MELTING  POINT:   268°  C.  (514°  F.). 

VAPORIZES  at  red  heat. 

ELECTRICAL  CONDUCTIVITY:   .013  compared  with  Ag. 

ALLOYS  with  most  metals.  The  alloys  with  Pb,  Sn,  Zn  and 
Cd  have  very  low  melting  points,  those  with  Cu  and 
Ni  are  very  hard. 


BISMUTH 


117 


CHEMICAL  BEHAVIOR:  Bismuth  is  very  resistant  toward  oxy- 
gen at  ordinary  temperatures,  but  is  easily  attacked  in  the 
fused  state,  though  not  so  readily  as  lead.  It  is  precipitated 
from  its  dissolved  or  fused  salts  by  lead.  It  is  dissolved  by 


FIG.  138  (x8). 

by  HC1  and  H2SO4  only  in  the  presence  of  oxidizing 
agents.  Hot  concentrated  H2SO4  dissolves  Bi  but  the  acid 
is  partly  reduced  to  SC>2.  In  all  compounds  of  technical 
importance  Bi  is  present  as  trivalent  cation. 


LEAD 

Sources 

Natural  Sources : 

GALENA  (PbS,  86.57%  pk) :  the  associated  minerals  and 
gangue  are:  metallic  sulphides  such  as  sphalerite,  pyrite, 
argentite  and  the  sulpho-salts  of  arsenic  and  antimony;  car- 
bonates, such  as  cerrusite,  smithsonite,  limestone  and  dolo- 
mite; further,  hematite  and  also  sandstone.  The  silver 
content  of  galena  varies  between  o.oi  and  i%;  in  the 
Rhenish  provinces  o.oi  to  .015%;  in  the  Hartz  0.05-0.1%. 

CERRUSITE  (PbCO3,    77.52%  Pb)  occurs  as   an   alteration 
product  of  galena.    It  is  found  in  the  upper  parts  of  galena 
deposits  and  is  associated  with  the  same  gangue. 
Other  Sources: 

LITHARGE:  PbO  from  cupellation. 

HEARTH  MATERIAL  (clay,  marl,  cement  and  less  often  bone 
ash)  saturated  with  litharge  from  cupellation. 

DROSS,  the  first  oxidation  product  from  lead  refining  or  cupel- 
lation, a  mixture  of  lead  oxide  and  lead  antimoniate. 

LEAD  MATTE,  an  intermediate  product  from  smelting  lead 
ores  after  the  roast-reduction  and  precipitation  processes; 
contains  lead  sulphide,  iron  sulphide,  copper  sul- 
phide, etc. 

Slags  which  contain  an  appreciable  amount  of  lead  are 
also  returned  to  the  smelter. 

ALLOYS  of  lead  with  zinc,  copper,  bismuth,  gold  and  silver. 

(A)   Concentration  Processes 

Mechanical,  Especially  Wet  Concentration,  is  used  for  ores 
carrying  galena.  Because  of  the  high  specific  gravity  of  galena, 
it  is  easy  to  concentrate  such  ores  up  to  a  lead  content  of  70-80%, 

118 


LEAD  119 

(B)   Extraction  of  Lead 

Roast-reaction  Process.  By  this  method  of  working,  sulphide 
lead  ores  are  said  to  be  treated  in  such  a  way  that  a  part 
of  the  sulphide  is  converted  by  oxidation  into  oxide  and  sul- 
phate with  which  the  unchanged  sulphide  reacts  to  form  SO2 
and  Pb.  The  chemical  reactions  taking  place  in  the  roasting 
and  the  reaction-smelting  are  explained  by  the  following  chem- 
ical equations: 
Roasting  : 

PbS  +  30  =  PbO  +  SO2      PbO  +  SO2  +  O  =  PbSO4. 

If  lead  carbonate  is  also  present  in  the  ores,  it  is  decom- 
posed as  follows  : 


=  PbOfC02. 
Reaction  Smelting: 

2PbO  +  PbS  -  3?b  +  SO2      PbSO4  +  PbS  =  Pb2  +  2SO2. 

If  the  unchanged  sulphide  and  the  oxidation  products 
are  not  present  in  the  above  proportions,  an  excess  of  oxide 
or  sulphide,  as  formed  by  the  first  equation,  will  remain 
unchanged.  The  excess  of  sulphate,  however,  works  as 
follows  : 


3PbSO4  +  PbS  =  4PbO  -f  4SO2. 

If  carbonate  is  still   present  during  the  reaction-smelting 
it  will  be  changed  as  follows: 


Schenck    and    Rassbach    have    determined    experimentally 


120 


METALLURGY 


the  conditions  of  equilibrium  between  Pb,  S  and  O.    Taking 
into  consideration  the  SO2  pressures,  the   conditions   given 
i  ii  in  m  Fig.  139  were  obtained  for 

the  formation  of  PbSO4,  PbO 
and  Pb. 

Field  I  is  the  formation  zone 
for  PbSO4;  here  the  princi- 
pal reaction  is  Pb2  +  2SO2  = 
PbSO4  +  PbS;  even  PbO  is 
converted  into  PbSO4  in  this 
zone: 


300 


100 

80 

28 

20 


700 


600 


500 


600         700         800 

FIG.  139. 


900°  C. 


Field  II  is  the  formation  zone 
for  PbO,  although  the  forma- 
tion of  PbSO4  is  not  yet 
impossible.  Between  PbO 
and  PbSO4,  solutions  and 
chemical  compounds  are 


formed.     The  main  reactions  here  are: 


and  then  the  following  equation, 
PbS  +  3PbSO4  = 


The  exact  boundaries  of  the  zones  in  field  II,  in  which  on 
the  one  hand  PbO  and  PbSO4  form,  and   on  the  other 
PbO  alone,  have  not  yet  been  determined. 
Field  III  is  the  zone  in  which  Pb  alone  forms  from  PbSO4, 
PbS  and  PbO: 


LEAD  121 

The  roast-reaction  process  is  applicable  only  to  ores  rich  in 
lead  with  not  over  4%  SiO2.  Various  modifications  in  the 
methods  used  have  sprung  up  because  of  local  conditions, 
especially  with  regard  to  wages  and  cost  of  fuel.  These  processes 
are: 

CARINTHIAN  PROCESS.  This  is  carried  out  in  small  rever- 
beratories  and  therefore  with  small  charges.  A  low  tem- 
perature is  maintained  and  the  roast  and  reaction  processes 
are  separate.  The  advantages  of  this  method  are:  low 
lead  loss  by  volatilization;  little  residue;  pure  lead.  Its 
disadvantages:  high  fuel  consumption  and  high  cost  for 
labor. 

ENGLISH  PROCESS.  Here  large  reverberatories,  large  charges, 
and  high  temperatures  are  used  from  the  start.  The 
fuel  consumption  and  cost  of  labor  are  low  compared 
with  the  Carinthian  method,  but  there  is  a  high  loss  by 
volatilization,  requiring  expensive  condensation  devices. 
There  also  remains  a  large  amount  of  lead-rich  residue. 

TARNOWITZ  PROCESS.  This  is  the  Carinthian  method  car- 
ried out  in  large  furnaces.  There  is  small  loss  by  volatili- 
zation, low  fuel  consumption  and  moderate  cost  for  labor. 
It  yields  a  pure  lead,  but  leaves  a  large  residue  in  rich  lead 
which  must  be  treated  in  special  furnaces. 

FRENCH  OR  BRITTANY  PROCESS.  The  work  is  done  in  large 
furnaces,  large  charges  and  a  long  roasting  period,  whereby 
considerable  lead  oxide  and  lead  sulphate  are  formed, 
which  necessitates  the  addition  of  fine  coal  during  the 
reaction  smelting. 

This  process  has  no  advantages  over  those  mentioned 
above,  but  on  the  contrary  a  high  loss  of  lead  is  entailed 
and  the  life  of  the  furnace  is  short. 

HEARTH  PROCESS.  The  ore,  with  fluxes  and  fuel  (charcoal), 
floats  upon  the  lead  in  a  crucible-like  hearth,  three  walls 
of  which  are  made  high.  Through  the  back  wall,  air  is 
blown  into  the  charge  above  the  molten  lead. 

The  plant  is  simple  and  quickly  constructed,  but  the 


122  METALLURGY 

process  entails  a  high  loss  of  lead  by  volatilization,  and  the 
laborers  are  endangered  by  lead  poisoning. 
Roast-reduction   Process.     By   this  method  the  ore,  after  a 

long-continued  roast,  is  melted  with  reducing  agents  to  produce 

lead. 

Roasting.  Qualitatively,  the  reactions  are  the  same  as  during 
the  roasting  period  of  the  roast-reaction  process.  Quantita- 
tively, however,  the  products  are  quite  different,  as  it  is  not 
intended  here  to  leave  sulphur  compounds  in  the  roasted  prod- 
uct for  the  purpose  of  reducing  the  oxides  that  are  produced 
in  the  roast.  The  sulphides  are,  therefore,  roasted  down  to 
the  smallest  amount  necessary  for  forming  matte  (i.e.,  if 
the  ores  contain  copper).  In  the  absence  of  metals  which  are 
to  be  concentrated  into  matte,  sulphates  also  are  not  desired 
in  the  roasted  ore.  The  sulphates  resulting  from  the  roast  are 
reduced  during  the  reducing  smelt.  This  re-formation  of 
sulphides  is  at  least  superfluous;  in  the  older  processes  the 
sulphate  was  decomposed  by  silica  at  the  close  of  the  roasting: 

2PbSO4  +  SiO2  =  Pb2SiO4  +  2SO2  +  O2, 

and  the  temperature  was  raised  correspondingly.  Apparatus: 
reverberatory  furnaces  with  hearths  narrowing  near  the  fire 
bridge. 

The  newer  processes  were  suggested  by  the  method  of 
Huntington  and  Heberlein,  in  which  a  mixture  of  galena  and 
lime  is  first  roasted  at  about  700°  C.  and  the  roast  product 
after  being  cooled  to  about  500°  is  blown  to  PbO  in  a  converter. 
Savelsberg  dispensed  with  the  preliminary  roasting  by 
moistening  well  the  mixture  of  ore  and  lime  and  blowing 
it  directly  in  a  converter. 

Finally  Carmichael  and  Bradford  treated  a  mixture  of  ore 
and  gypsum  directly  in  a  converter. 

In  all  three  cases,  the  attempt  is  made  to  fulfil  the  con- 
ditions explained  by  Schenck  and  Rassbach  for  obtaining 
oxides,  either  directly  or  with  the  intermediate  formation  of 


LEAD  123 

lead  sulphate,  and  to  bring  the  ore  to  a  proper  condition 
for  treatment  in  the  subsequent  reducing  smelt.  Plumb- 
ates  form  at  the  beginning  as  well  as  at  the  closing  stages 
of  these  treatments. 

The  discovery  of  Doeltz  that   CaSO4  does  not  react 

with  PbS  does  not  affect  the  Carmichael-Bradford  process. 

In  the  plant  operated  in  Australia  the  reacting  substances 

are  not  CaSO4  +  PbS,  but  CaSO4  +  SiO2  +  PbS. 

Reduction  Smelting.     In  the  roast-reaction  process,  the  unde- 

composed  PbS  serves  as  a  reducing  agent  for  the  oxides  and 

sulphates  formed  in  the  roast.     In  this  process,  however, 

the  reduction  is  effected  by  carbon  and  carbon  monoxide: 

2PbO  +  C  =  Pb2  +  CO2.    PbO  +  CO  =  Pb  +  CO2. 

The  direct  reduction  of  lead  silicate  in  the  original  ore, 
or  of  that  produced  in  roasting,  cannot  be  effected  by  carbon; 
the  lead  oxide,  therefore,  in  order  to  be  acted  upon  by  the 
reducing  agents  must  be  set  free  from  the  silicate  by  fluxes 
which  also  form  an  easily -fusible  slag: 

Pb2SiO4  +  CaO  +  FeO  =  CaFeSiO4  +  2PbO. 

The  roast  reduction  process  is  applicable  to  nearly  all 
lead  ores;  if  the  silica  is  low,  a  silicious  flux  may  be  required. 
In  contrast  to  the  roast-reaction  process,  in  which  only 
one  apparatus  is  used  for  both  stages  of  the  process,  the 
roast  reduction  process  requires  two  forms  of  apparatus. 
Usually  the 

ROASTING  is  carried  on  in  reverberatory  furnaces  or  in 
funnel-shaped  or  bowl-shaped  converters  having  tuyeres 
at  the  bottom. 

THE  REDUCING  SMELT  is  always  carried  out  in  shaft  fur- 
naces. 

The  Precipitation  Process.  In  this  process  the  roasting  of  the 
sulphide  ores  as  such  is  avoided,  because  the  roasting  is  always 
difficult  to  complete  on  account  of  the  fusibility  of  galena,  lead 
oxide,  and  the  other  lead  compounds  that  result  from  the 


124  METALLURGY 

roast.    The   sulphide   ores,  therefore,  are  fused  directly  with 
fluxes  containing  metallic  iron  : 

Pb  +  FeS. 


The  iron  sulphide  resulting  from  this  reaction  exerts  a  strong 
solvent  action  on  the  other  sulphides.  If  this  fact  were  not 
taken  into  consideration,  the  first  iron  sulphide  produced  in 
the  reaction  would  unite  with  the  galena  and  thus  prevent  its 
decomposition  by  iron.  Hence  an  excess  of  galena  is  used  over 
the  quantity  required  by  the  above  equation: 

#PbS  +  Fe  =  FeS(PbS).r_i+Pb 

This  production  of  lead  matte  has  several  advantages: 

THE  MATTE  may  be  roasted  in  low  shaft  furnaces,  producing 
concentrated  gases  which  may  be  utilized  for  making 
sulphuric  acid.  This  advantage  is  not  possessed  by  the 
roasting  apparatus  used  in  the  roast-reaction  and  roast- 
reduction  processes.  With  the  roasted  lead  matte,  a  large 
part  of  the  iron  used  in  the  original  fusion  is  returned 
to  the  process,  as  this  product  consists  chiefly  of  lead 
oxide  and  iron  oxide: 

2  (PbS.FeS)  +  130  =  2PbO  +  Fe2O3  +  4SO2. 

The  precipitation  process,  because  of  the  above  facts, 
may  be  used  to  advantage  on  ores  which  do  not  contain 
large  quantities  of  other  sulphides. 

The  constitution  of  lead  matte  has  been  cleared  up  by 
the  investigations  of  H.  Weidtmann,  who  found  that  no 
chemical  compounds  exist  between  PbS  and  FeS,  but 
a  eutectic  exists  having  the  approximate  composition 
PbS  +  FeS  and  a  melting  point  of  780°  C.  (1435°  F.). 
Apparatus  for  the  Roast-reduction  and  Precipitation  Proc- 

esses : 

Roasting  Apparatus: 

HEAP  ROASTING  is  used  as  a  preliminary  operation  for  ores 
high  in  zinc  and  pyrite,  the  object  being  to  sulphatize  and 
leach  out  the  zinc. 


LEAD  125 

KILNS  for  roasting  matte  (see  Copper;. 

REVERBERATORY  FURNACES.  In  the  early  methods ,  the 
long-hearth,  hand  reverberatory  was  used  almost  univer- 
sally. The  hearths  were  often  up  to  85  ft.  in  length  and 
from  8  to  1 6  ft.  in  width.  If  floor  space  was  lacking,  furnaces 
were  constructed  having  two  hearths,  one  above  the  other, 
each  from  40  to  50  ft.  long.  At  the  sides  of  the  hearths  were 
working  doors,  5  to  6J  ft.  apart.  The  reason  for  build- 
ing such  long  hearths  was  because  of  the  nature  of  the  raw 
material  and  of  the  roasted  products.  They  are  readily 
fusible  and  require,  therefore,  a  low  temperature  and  a 
correspondingly  long  time  for  the  roasting.  In  spite  of 
the  endeavor  to  pass  the  ore  quickly  through  the  furnace, 
a  considerable  quantity  remains  continually  on  the  hearth, 
and  is  moved  from  time  to  time  in  small  parcels  toward 
the  firebox,  where  the  final  roasting  takes  place. 

The  construction  of  the  hearth  depends  upon  the  special 
object  of  the  roast. 

SLAG  ROASTING.  By  this  method  the  reaction  between 
PbSO4  and  SiC>2,  mentioned  above,  is  carried  out  as  com- 
pletely as  possible.  The  portion  of  hearth  lying  nearest 
the  firebox  is  constructed  in  such  a  way  and  of  such  dimen- 
sions that  the  roasted  product  is  fused  here.  The  slagging 
hearth,  therefore,  is  made  narrower  and  lower  at  this  point 
in  order  to  concentrate  the  heat  and  collect  the  fused 
material.  At  the  side  of  the  fusion  hearth  farthest  from 
the  firebox,  the  hearth  widens  and  usually  becomes  higher 
in  order  that  the  temperature  may  be  greatly  lowered 
for  the  roasting  (Figs.  140-141). 

SINTER-ROASTING.  In  this  method  a  complete  desul- 
phurization  is  not  sought;  the  chief  aim  is  to  heat  the  pul- 
verulent roasted  product  only  enough  to  sinter  it.  The 
furnaces  required  for  this  work  have  a  short  sinter-hearth 
near  the  fire-bridge,  to  which  is  joined  the  roasting-hearth 
proper  a  step  higher,  making,  as  it  were,  a  terrace.  The 
sinter-hearth  does  not  need  to  be  narrowed. 


126  METALLURGY 

NON-SINTER  ROASTING.  In  this  work  the  product  of 
roasting  remains  in  a  pulverulent  state.  The  hearth  of 
the  furnace  used  for  this  method  is  a  slightly  inclined 
plane,  sloping  away  from  the  flue-bridge  toward  the 
fire-bridge. 
FOR  REVERBERATORY  FURNACES  WITH  STATIONARY 

HEARTHS  many  mechanical  stirring  devices  have  been 

used.     These  have  been  described  under  Copper,    pp. 

63  to  71. 
REVERBERATORY    FURNACES    WITH    MOVABLE    HEARTHS 

are  used  in  the    Huntington-Heberlein    process,   espe- 


FIGS.  140-141. — Reverberatory  Furnace  with  Slagging  Hearth. 

cially    those    with    revolving    circular    hearths,    having 

stationary  rabbles,  similar  to  the  heating  furnaces  used 

in  the  manufacture  of  coal  briquettes. 
CONVERTERS  are  used  in  the  following  processes: 

HUNTINGTON-HEBERLEIN:  conical,  cast-iron  vessels  5-8 
feet  wide  at  the  top  and  5-6^  ft.  deep,  fitted  at  the 
bottom  with  grates  to  support  the  charge  and  to  distribute 
the  blast,  which  is  also  admitted  at  the  bottom.  Capacity 
5-8  tons. 

CARMICHAEL-BRADFORD:  conical,  cast  iron  vessels; 
upper  diameter  6  ft.,  lower  diameter  4  ft.,  depth  5  ft. 


LEAD 


Above  the  bottom,  which  is  arched  downward,  is  placed 
a  perforated  shell  arched  upward  which  acts  as  a  grate. 
The  blast  is  admitted  into  the  intermediate  space  thus 
formed.  Capacity,  4  tons  (Fig.  142). 

SAVELSBERG.  Nearly  hemispherical,  cast-iron  pots  about 
6 1  feet  in  diameter.  Air  admitted  from  below.  Capacity 
8  tons.  Blast  required,  250  cu.ft.  per  min.  Blast  pressure 
at  start,  4  to  8  in.  of  water,  later  20  to  24  in.  Time 
of  blowing,  18  hours  (Fig.  143). 
Apparatus  for  the  Reduction  and  Smelting : 

SHAFT    FURNACES.      Of    the    older  constructions  the    low 


FIG.  142. — Carmichael-Bradford 
Converter.     Scale,  i  :  50. 


FIG.  143. — Savelsberg  Converter. 
Scale,  i :  50. 


furnaces  (Krummofen)  have  almost  wholly  disappeared, 
while  the  medium  furnaces  (Halbhochof  en)  are  found  only 
in  a  few  plants.  The  high  furnaces  (Hochofen)  are  used 
almost  exclusively  in  modern  plants.  They  are,  of  course, 
much  smaller  than  the  blast  furnaces  used  in  the  iron 
industry. 

The  "  Krummofen "  are  low  shaft  furnaces  not  more 
than  6J  ft.  in  height,  with  square,  rectangular  or  trapezoidal 
cross-section.  They  usually  have  only  one  tuyere  which  is 
placed  in  the  back  wall.  Various  forms  of  these  furnaces 


128 


METALLURGY 


are  described  in  the  oldest  metallurgical  literature  ("Agric- 
ola  de  rebus  nietatticis,"  1657,  pp.  313-319;  Schluter, 
"  Huttenwerke,"  1738,  tables  XX-XXX). 

"  HALBHOCHOFEN."  These  are  low  shaft  furnaces,  7-14 
ft.  in  height  from  the  tuyere  level  to  the  throat.  They  are 
usually  of  trapezoidal  cross-section  with  from  2  to  8  tuyeres 
on  one  side  of  the  trapezoid.  Of  these  furnaces  the 
Stolberg  furnace  has  been  most  widely  adopted  and  was 


FIG.  144. — Pilz-Freiberg.       FIG.  145. — Upper  Harz.     FIG.  146. — Allis-Chalmers. 

also  used  until  recently  in  Freiburg.  They  are  still  in 
operation  at  the  lead  plants  in  Stolberg  and  Mechernich. 

"  HOCHOFEN."  Although  the  name  is  usually  applied  to 
blast  furnaces  having  at  least  13  ft.  between  the  tuyere 
level  and  throat,  shorter  furnaces  of  this  type  are  found 
in  some  lead  plants. 

Modern  blast  furnaces  are  built  with  circular  and  rect- 
angular horizontal  cross-sections  (Figs.  144-147).  The 
shaft  is  usually  made  of  brick  and  the  bosh  and  smelting 
zone  are  water-jacketed.  The  throat  is  open,  except  during 
blowing  in  and  blowing  out,  when  it  is  covered  with  a 


LEAD 


129 


hood.     Downcomers   start  from  the  centre  of  the  throat 
or  from  the  side  under  the  feed  floor.     They  are  usually 


FlG.  147. — North  American  Water-jacket  Furnace.     Scale,  i  :  150. 

made  of  encased  sheet-iron  pipes.     Fore-hearth:   crucible 
furnace  or  Arent's  siphon  tap. 

The  direct  electric  smelting  of  lead  ores  has  not  yet  been 
successful. 

(C)    Lead  Refining 

The  crude  lead  obtained  in  the  preceding  processes  may  con- 
tain various  quantities  of  all  the  metals  found  in  the  ores,  also 
metallic  compounds  such  as  sulphides,  arsenides,  antimonides, 
etc.  Among  the  impurities  are:  Au,  Ag,  Cu,  Sn,  Sb,  As,  Bi,  Co, 


130  METALLURGY 

Ni,  Fe,  Zn,  and  S.     They  may  be  separated  from  the  lead  by  the 
following  operations : 

Liquation  and   Crystallization  Processes  (see   Silver).     In  this 
way  Au,  Ag,  and  Cu  are  removed  by  being  concentrated  in 
a  part  of  the  lead. 
Oxidation  of  the  Impurities  (true  lead  refining). 

(a)  Oxidation  of  S,  As,  Sb,  Bi  by  atmospheric  oxygen. 

(b)  Oxidation  of  Zn,  Fe,  Co,  and  Ni  by  steam. 

Both  of  the  above  methods  have  already  been  described 

under  Silver. 
Oxidation   of  the  Lead,  the   so-called    cupellation   process  for 

recovering  precious  metals  (see  Silver  and  Bismuth). 
Electrolysis.     After  several   unsuccessful  attempts   by  Keith, 

Tommasi  and  others,  this  process  was  successfully  applied 

by  Betts,  in  1907,  to  the  refining  of  antimonial  lead  and  base 

bullion.     A  description  and  sketches  of  a  complete  plant 

are  given  in  Metallurgie,  1908,  5,  68. 

ANODES:  base  bullion. 

CATHODES:  pure  lead. 

ELECTROLYTE  :  aqueous  solution  of  silicon  fluorides  containing 
22.7  oz.  PbSiF6 
9.4  oz.  H2SiF6 
6.7  oz.  gelatin  per  100  gallons. 

TEMPERATURE:  between  17°  and  57°  C.  (63  and  135°  F.) 

CURRENT  DENSITY:  10  to  12  amp.  per  square  foot  of 
cathode. 

POTENTIAL:  0.15-0.36   volt. 
Properties  of  Lead: 

SPECIFIC  GRAVITY:  11.4. 

COLOR:  bluish  gray,  shiny. 

MECHANICAL  PROPERTIES:  low  tensile  strength,  high  duc- 
tility. 

STRUCTURE.  Ingot  fracture  (Fig.  148).  Crystallized  shell 
from  which  the  molten  metal  was  removed  before  freezing 
was  completed  (Fig.  149).  Granular  structure  obtained 
by  etching  the  bottom  surface  of  a  lead  ingot  (Fig.  150). 


LEAD 


131 


MELTING  POINT:  327°  C.  (621°  F.) 

VAPORIZATION  takes  place  at  red  heat,  though  the  boiling 
point  is  between  1200  and  1300°  C.  (2200  and  2375°  F.) 


FIG.  148. 


FIG.  149. 

THERMAL  AND  ELECTRICAL  CONDUCTIVITY:  low,  the  latter 

0.0756  that  of  Ag. 
ALLOYS  with  most  metals.     Its  most  important  metallurgical 

alloys  are  those  with  the  precious  metals,  for  which  it  is 


132  METALLURGY 

utilized  as  a  solvent:  with  zinc  (limited  solubility),  with 
antimony  (hard  lead)  and  with  tin  (solder). 
CHEMICAL  BEHAVIOR:  readily  oxidized  on  the  surface  by 
O,  H2O,  and  CO2  of  the  atmosphere.  The  coating  of 
oxide  and  basic  carbonate  thus  formed  prevents  further 
corrosion. 

In  the  same  way  the  action  of  HC1  and  H2SO4  is  restricted 
even  in  the  presence  of  oxygen.     Without  the  presence 


FIG.  150  (x  33)- 

of  oxygen  or  oxidizing  agents  in  the  acid,  it  is  hard  to 
dissolve  lead  in  acids,  because  the  electrolytic  potential 
is  only  +.151  against  H. 

There  are  two  common  oxides,  PbO  and  PbO2,  the 
former  basic,  the  latter  acid.  These  unite  to  form  a 
plumbate,  Pb2PbO4  (which,  together  with  free  PbO, 
constitutes  red  lead). 

Of  free  sulphides,  only  PbS  is  known;  it  dissolves 
readily  in  other  sulphides  (matte),  but  forms  few  chemical 
compounds. 


TIN 

Sources 

Natural  Sources : 
OXIDIZED  ORES. 

Tin-stone,  or  cassiterite,  SnO2.  In  primary  deposits,  lode 
tin;  in  secondary  deposits,  stream  tin.  Gangue:  acid 
eruptive  rocks,  particularly  granite,  also  quartz,  CaF2, 
FeWO4,  CaWO4,  sulphides  such  as  Cu2S,  PbS,  Bi2S3, 
MoS3,  etc.,  and  arsenides  such  as  FeAsS. 
SULPHIDE  ORES. 

Tin    pyrites,  FeCu2.SnS4,  which  is  much  rarer  than  tin- 
stone. 

Other  Sources : 

METALLIC  INTERMEDIATE  PRODUCTS  AND  BY-PRODUCTS: 
Hard  Head  and  Liquation  Dross  (Fe-Sn  alloy). 
Tin  plate  clippings  (Fe  covered  with  a  coating  of  Sn). 
White  Metal  in  the  form  of  impure  turnings:  Sn,  Pb,  Cu, 

Zn,  Sb. 

OXIDIZED  BY-PRODUCTS: 
Slags:  Silicates  and  Stannates. 
Ashes:  Oxides  mixed  with  metallic  grains. 

(A)   Concentration 

Mechanical  Concentration.  On  account  of  the  high  specific 
gravity  of  tin-stone,  viz.,  6.8  to  7,  the  mechanical  concentration 
of  tin  ore  is  carried  out  quite  commonly,  and  sometimes  by 
the  simplest  means. 

Chemical  Concentration.  Among  the  minerals  accompanying 
tin-stone  are  found,  as  mentioned  above,  sulphides,  arsenides 
and  tungstates.  The  first  two  are  almost  always  present,  but  of 

133 


134  METALLURGY 

the  last  mentioned  only  a  few  ores  contain  enough  to  cause 

trouble. 

The  removal  of  S  and  As  is  effected  by  an  oxidizing  roast  in  a 
simple  roasting  furnace,  usually  fed  by  hand.  For  ores 
with  much  sulphide  of  Fe  and  Cu,  it  has  been  pro- 
posed to  carry  out  the  roasting  in  such  a  way  that  the  sul- 
phates of  these  metals  are  formed,  which  can  then  be  leached 
out,  but  such  a  process  does  not  seem  to  have  met  with  much 
practical  application. 

The  removal  of  tungstates  is  necessary  with  ores  containing 
much  tungsten,  because  tungstates  favor  the  passage  of  tin 
into  the  slag.  Even  a  small  percentage  of  tungstate  suffices 
not  only  to  carry  over  considerable  amounts  of  tin  into  the 
slag  on  smelting  the  ore,  but  also  hinders  the  reduction  of 
tin  in  the  subsequent  smelting  of  the  slag.  Hence,  even  in 
smelters  that  work  without  previous  concentration,  with 
ores  containing  small  amounts  of  tungsten,  it  is  worth  while 
to  take  the  slags  in  which  considerable  tin  has  accumulated, 
and  remove  the  tungsten  in  order  to  recover  the  tin  by  another 
smelting.  The  removal  of  tungsten  is  effected  by  an  oxidizing 
roast  of  the  pulverized  ore  or  slag,  with  sufficient  Na2CC>3 
to  convert  the  W  into  Na2WO4.  The  latter  is  removed  by 
lixiviation,  whereupon  the  tin  can  be  obtained  without  diffi- 
culty. Fuller  details  concerning  such  work  will  be  given 
under  Tungsten. 

(B)    Extraction 

Inasmuch  as  tin-stone  is  the  principal  raw  material,  which, 
when  necessary,  is  concentrated  or  purified  by  the  above-men- 
tioned work,  the  chief  operation  of  the  tin  smelter  is  of  a  com- 
paratively simple  nature.  It  consists  of  a  reducing  fusion 
of  the  ore,  and  other  oxidized  waste  products,  and  the  subsequent 
working  over  of  the  slag.  During  the  last  two  decades  it  has 
also  been  found  possible  by  an  electrolytic  process  to  recover 
tin  from  the  scrap  obtained  in  the  preparation  of  tin  plate. 


TIN  135 

The    Smelting    Operations    for    Oxidized    Ores    and   Waste 
Products  are  the  following. 

i.  Reducing  Fusion.  Concerning  the  reactions  that  take 
place  during  this  process,  singular  views  are  to  be  found 
in  metallurgical  literature.  The  ready  reducibility  of  SnO2 
by  KCN,  which  is  preferred  as  a  flux  and  reducing  agent  in 
making  the  blow-pipe  test  for  tin,  has  led  to  the  assumption 
that  during  smelting  the  carbon  of  the  coal  used  for  effecting 
the  reduction  is  partly  converted  into  cyanide,  which  then 
acts  as  in  the  blow-pipe  test.  There  has  never  been  any 
proof  of  this  and,  moreover,  the  assumption  is  very  improbable, 
because,  in  most  cases,  the  slag  is  not  kept  basic  enough 
nor  hot  enough  for  the  formation  of  cyanide.  An  experimental 
investigation  by  Mattonet,  in  the  laboratory  of  the  author, 
has  shown  that  the  formation  of  a  singulo-silicate  slag  is 
most  favorable  for  the  extraction  of  the  tin  and  for  the  slag- 
ging off  of  the  other  metals. 

According  to  the  nature  of  the  ore,  the  fuel,  and  the  other 
working  conditions,  it  has  been  customary  to  make  use  of 
either  a  blast  furnace  or  a  reverberatory  furnace  for  smelting. 
The  above-mentioned  researches  of  Mattonet,  however, 
have  established  the  fact  that  an  electric  furnace  can  be  of 
sendee  for  producing  slags  low  in  tin.  In  both  blast-furnace 
and  reverberatory  smelting,  the  extraction  of  the  tin  is  made 
difficult  by  the  unfavorable  position  of  the  reduction  temper- 
ature (1000°  to  1100°  C.)  in  respect  to  the  melting  point  (232°) 
and  the  boiling-point  of  the  metal.  The  latter,  to  be  sure, 
is  said  to  be  above  2100°,  but  volatilization,  of  the  metal 
takes  place  at  temperatures  far  below  the  boiling  point. 
The  furnace  charge  must,  therefore,  be  heated  to  1100°,  and 
the  combustion  gases,  which  furnish  the  source  of  the  heat 
and  its  transference,  must  be  kept  at  least  200°  hotter,  i.e., 
at  about  1300°,  so  that  the  Sn  is  heated  to  about  1000°  above 
its  melting-point,  at  which  temperature  it  shows  a  tendency 
to  volatilize  with  the  reaction-gases  and  also  possesses  a  high 
solution  tension  which  favors  its  passing  into  the  slag.  In 


136  METALLURGY 

the  electric  process,  particularly  since  the  charge  itself  can 

take   part   in  the  heat   transference,   the  surroundings   are 

kept  colder  than  the  charge. 

THE  BLAST-FURNACE  PROCESS,  which  is  the  oldest  method 
of  working,  is  applicable  if  a  very  pure,  lumpy  ore, 
and  a  pure  fuel,  e.g.,  wood  charcoal,  are  available.  The 
smelting  takes  place  sometimes  without  any  additions 
of  flux,  and  sometimes  with  slags  obtained  from  tin  ores, 
used  in  amounts  varying  from  25-50%  of  the  weight  of  ore. 
In  the  Chinese  and  Malay  tin  districts,  the  work,  in  some 
cases,  is  still  carried  out  with  very  primitive  apparatus. 
Whereas  the  original  blast-furnaces  of  the  Sunda  Islands 
consisted  of  pits,  about  20  in.  deep  and  12  to  16  in. 
wide,  dug  in  the  earth  and  provided  with  hand-driven 
blast  which  passed  through  hollow  tree  trunks,  the  Chinese 
draft  and  blast  furnaces  are  prepared  by  tamping 
clay  within  a  barrel-shaped  form  made  of  bamboo 
rods  and  then  carving  out  the  shaft  and  the  holes  for 
blowing  and  tapping.  These  furnaces  are  built  up  to 
6  ft.  in  height,  with  a  shaft  up  to  16  in.  wide  and  about 
5  ft.  3  in.  high.  Such  furnaces,  when  used  for  smelting 
the  ore,  will  work  with  natural  draft,  provided  the  mate- 
rial is  sufficiently  coarse;  and  those  used  for  smelting  the 
slag  work  with  a  blast,  which,  as  in  the  above  instance,  is 
produced  by  hand,  using  hollow  tree  trunks  that  are  pro- 
vided with  wooden  pistons.  (Fig.  151.)  Even  the  larger 
blast  furnaces,  such  as  are  used  in  India  at  works  con- 
ducted by  Europeans,  as  well  as  in  Bohemia  and  in  Alten- 
berg,  are  built  so  short  and  narrow  that  they  can  almost 
be  classed  as  "  dwarf  furnaces."  Concerning  the  con- 
struction of  such  furnaces,  enough  was  given  under 
Copper  and  Lead,  so  that  it  is  unnecessary  to  go  into 
further  details  here. 

The  method  of  smelting  tin  ores  which  is  in  common  use 
to-day  is  by  means  of  the  REVERBERATORY  FURNACE, 
because  the  ore  is  usually  fine  grained  in  nature  and  of 


TIN 


137 


low  grade,  so  that  the  use  of  the  blast  furnace  is  out  of  the 
question.  Although  formerly  considerable  importance  was 
attached  to  the  nature  of  the  fuel,  inasmuch  as  only  those 
varieties  of  coal  are  suitable  which  produce  long  flames, 
to-day  this  difficulty  can  be  overcome  by  the  use  of  gas 
firing.  Even  in  this  case  the  furnace  is  almost  always 
charged  merely  with  ore  and  coal;  when  the  gangue  is  too 
acid  it  is  sometimes  necessary  to  add  limestone  and  small 


FIG.  151. — Chinese  Furnace  for  Smelting 
Tin  Ore. 


amounts  of  fluorspar.  In  order  to  prevent  mechanical 
losses  of  the  powder,  the  mixture  is  usually  moistened 
with  water.  After  being  introduced  into  the  hot  furnace, 
the  charge  is  levelled  off  and  heated  strongly  for  from  5 
to  8  hours  with  closed  doors  and  nearly  checked  draft; 
after  the  mass  has  softened,  it  is  strongly  heated  for 
from  45  minutes  to  an  hour,  the  metal  is  tapped  and  the 
slag,  which  solidifies  readily  on  cooling,  drawn  off.  In 
other  works,  the  slag  is  thickened  by '  working  in  some 


138  METALLURGY 

coal  so  that  it  can  be  removed  through  the  slag  door  before 
tapping  the  metal,  which  is  then  covered  with  only  little 
slag.  According  to  the  size  of  the  reverberatory  furnace, 
the  amount  of  ore  used  in  a  single  charge  varies  from 
1600  Ibs.  to  4  tons,  the  time  of  heating  from  6  to  12  hours, 
the  consumption  of  fuel  from  60  to  120%  of  the  ore,  the 
labor  employed  from  i  to  3  men  per  furnace,  and  the 
smelting  losses  up  to  7%  Sn. 

As  regards  the  construction  of  the  reverberatory  furnaces, 
the  following  points,  which  were  partly  discussed  under 
Bismuth,  should  be  considered:  Sn  penetrates  readily 
into  the  narrowest  fissures  of  masonry,  and  since  it  Is 
practically  impossible  to  make  an  absolutely  impenetrable 
hearth,  this  is  not  built  any  stronger  at  the  bottom  than 
is  absolutely  necessary  for  holding  the  heat  within  the 
smelting  zone.  In  all  the  newer  constructions  of  rever- 
beratory furnaces,  the  hearth  is  built  upon  rails,  or  other 
supports,  independent  of  the  furnace-masonry,  so  that  it 
can  be  easily  repaired  without  destroying  the  latter, 
and  beneath  the  hearth  is  an  air  space  which  is  well 
ventilated  and  in  some  cases  even  kept  filled  with  water. 
The  reason  for  keeping  this  space  cold  either  by  ventila- 
tion or  by  water,  is  to  cause  the  immediate  solidification 
of  any  tin  that  trickles  through.  On  account  of  the 
large  dimensions  of  the  furnaces,  it  is  not  feasible  to 
make  movable  hearths,  as  in  the  bismuth  and  antimony 
industries  (cf.  p.  114,  Figs.  134-137).  It  is  necessary, 
however,  to  have  the  fireplace  and  flue  built  in  such  a 
way  that  they  are  separated  by  air  spaces,  or  some  insulat- 
ing material,  from  the  parts  of  the  furnace  containing  the 
hearth.  Otherwise  the  molten  tin  would  tend  to  run 
toward  the  fire  bridge  or  flue  bridge  and  find  a  particularly 
favorable  opportunity  to  penetrate  through  them. 
The  smelting  of  tin  in  an  electric  furnace  has  been  carefully 
studied  by  Mattonet,  in  the  author's  laboratory.  Exper- 
iments performed  in  the  attempt  to  smelt  impure  tin  ores, 


TIN  139 

without  a  preliminary  roasting,  in  such  a  way  that  the  Cu 


I 

b 

o 

I 


and  a  part  of  the  Fe  would  be  converted  into  a  matte  and 
the   Pb   into   an   acid   slag,   proved   unsuccessful.     Even 


140  METALLURGY 

the  sulphur  in  the  ore  passed  to  a  considerable  extent  into 
the  crude  tin,  which  contained  up  to  2.7%  S.  The  yield 
of  Sn  under  these  conditions  was  also  unsatisfactory. 
Better  results  were  obtained  with  roasted  ores  and  an 
approximately  neutral  slag,  in  which,  however,  by  adding 
some  soda  to  the  charge  the  melting-point  was  lowered 
somewhat  and  Na2O  made  a  part  of  the  basic  constituents. 
Since  at  most  only  450  Ibs.  of  Na^COs  (cost  $2.50)  are 
needed  for  a  ton  of  Sn,  the  cost  of  working  is  not 
increased,  because  more  than  enough  Sn  is  prevented 
from  going  into  the  slag  to  compensate.  The  slag 
obtained  was  perfectly  free  from  Sn  globules  (in  the  case 
of  other  methods  of  smelting  it  is  permeated  by  pellets 
of  Sn),  and  contains  only  from  i  to  1.5%  of  slagged  Sn,  as 
compared  with  from  2  to  5%  by  other  processes.  Direct 
heating  by  resistance  is  applicable  for  this  method  of  smelt- 
ing on  account  of  the  relatively  good  conductivity  of  the 
charge  and  of  the  slag.  In  this  way,  the  proper  tempera- 
ture for  the  reduction  of  the  Sn  is  maintained  and  is  not 
much  exceeded  in  the  bath.  Since,  moreover,  the  blast, 
which  in  one  method  tends  to  prevent  the  flowing  together.of 
the  Sn  globules,  is  absent  and  an  acid  lining  of  the  hearth, 
which  causes  losses  by  slagging  and  by  volatilization,  is 
unnecessary,  the  direct  yield  of  Sn  is  so  good  that,  whereas 
it  is  absolutely  necessary  to  recover  Sn  from  the  slag  in 
smelting  by  the  blast  furnace  or  by  the  reverberatory 
furnace,  in  the  case  of  electric  smelting  this  subsequent 
treatment  of  the  slag  can  be  dispensed  with  altogether  or 
limited  at  the  most  to  one  smelting.  Consequently  the 
smelting  of  rich  and  poor  slags,  as  described  below  under 
2  and  3,  refers  only  to  blast-furnace  and  reverberatory 
furnace  practice. 

2.  Smelting  of  Rich  Slags.      The  slags  obtained  by  smelting 
the  ore  contain,  as  mentioned  above,  tin  which  is  mechan- 
ically enclosed,  as  well  as  that  which  is  slagged.     As  soon 
as   a   sufficient   amount   has   accumulated,    it   is  smelted, 


TIN  141 

usually  in  a  reverberatory  furnace,  together  with  other 
metallic  waste  (dross  from  the  refining,  and  the  "hard- 
head "  or  residue  which  remains  on  the  hearth),  and,  in  case 
not  enough  iron  is  present  to  cause  the  separation  of  the 
slagged  tin,  some  scrap  iron  is  added  also.  Obviously  it  is 
necessary  to  conduct  the  work  with  a  reducing  flame  and  with 
some  fuel  added  to  the  charge. 

Since  the  smelting  of  rich  slags  requires  a  higher  tem- 
perature than  that  of  ores,  this  process  is  often  carried  out 
for  the  purpose  of  seasoning  new  hearths  which  are  to  be  used 
for  smelting  ore.  A  considerable  part  of  the  slag  adheres 
to  the  masonry  of  the  hearth  and  either  remains  solid 
during  subsequent  smelting  of  ores,  or  else  is  so  viscous  that 
metal  does  not  flow  through  it. 

3.  Smelting  of  Poor  Slags.  This  is  carried  out  in  many  works 
in  blast  furnaces,  usually  in  small  water-jacketed  furnaces 
having  a  circular  cross-section;  large  pieces  of  coal  or  coke 
are  mixed  with  the  charge. 

Although  d  fairly  pure  tin  is  obtained  directly  by  the 
two  former  operations  (the  crude  metal  from  rich  slags  even 
contains  95%  Sn),  in  this  case  the  crude  metal  is  a  fairly 
rich  iron  alloy  (80%  of  Sn,  20%  of  Fe). 

Electrolytic  Solution  and  Deposition.  For  twenty  years  this 
process  has  been  carried  out  in  working  up  tin-plate  waste. 
The  object,  in  most  cases,  is  to  recover  the  tin  from  old 
scrap  made  in  the  manufacture  of  tin-plate.  According  to 
the  thickness  of  the  sheet  iron  which  was  given  a  coating  of 
Sn,  the  material  contains  from  3  to  5%  of  Sn.  As  a  rule,  how- 
ever, it  is  not  safe  to  reckon  on  more  than  2  or  3%.  The 
recovery  of  the  tin  has  been  carried  out  most  successfully  in 
the  factory  of  T.  Goldschmidt  at  Essen.  Most  other  works 
have  made  use  of  the  same  method  and  profited  by  the  experi- 
ence gained  there.  The  apparatus  and  conditions  are  as 
follows : 

ELECTROLYZING    VESSELS:     iron    tanks.     These    are    con- 
nected with  the  circuit  and  serve  as  cathodes. 


142 


METALLURGY 


ANODES:  the  tin-plate  clippings  packed  not  too  firmly  in 
a  basket  made  of  coarse  iron  wire  (about  100  Ibs.  of  clip- 
pings are  placed  in  a  basket). 

CATHODES:  sheet  iron  and  the  walls  of  the  tank. 

ELECTROLYTE:  this  consists  at  first  of  a  caustic  soda  solu- 
tion containing  at  the  most  9%  of  NaOH,  corresponding 
to  7%  Na2O.  During  the  electrolysis  the  caustic  soda  is 
partly  converted  into  carbonate  and  partly  into  stannate. 
Of  the  original  7%  of  Na2O,  3  to  3.5%  remains  as  the 
free  hydroxide  while  i  to  1.5%  of  Na2O  is  converted  into 
stannate  and  1.7  to  2.8%  of  Na2O  unites  with  CO2.  If 
this  amount  of  carbonate  is  exceeded,  the  electrolyte 


FIG.  155. 


FIG.  156. 


must  be  worked  over,  by  treating  with  Ca(OH)2,  whereby 
CaCO3  is  precipitated  and  NaOH  formed. 
TEMPERATURE:    70°  C.  (158°  F.). 
POTENTIAL:    1.5  volts. 

Concerning  the  reactions  that  take  place  during  the  elec- 
trolysis there  is  but  little  reliable  information.  It  is  known 
that  the  Sn  goes  into  solution  as  stannate,  and  therefore 
forms  a  complex  anion,  and  that  it  is  consequently  indirectly 
precipitated  (by  Na  ?) . 
PRODUCTS  : 

ELECTROLYTIC  TIN.  Inasmuch  as  old  tin-ware  that  is 
worked  up  in  this  way  contains  more  or  less  solder,  the 
Sn  obtained  by  electrolysis  usually  contains  some  Pb. 
For  this  reason,  the  Sn  is  deposited  in  the  form  of 


TIN  143 

crystalline  grains  and  does  not  form  coherent  plates. 
On  being  dried  it  becomes  covered  with  a  layer  of  oxide, 
so  that  it  must  be  subjected  to  a  reducing  smelt. 

LEAD  CARBONATE:  the  first  precipitation  product  in 
working  up  impure  electrolytes. 

STANNIC  ACID,  or  SnO2:  the  second  precipitation  prod- 
uct from  impure  electrolytes. 

(C)   Tin  Refining 

It  is  clear,  from  what  has  been  said  concerning  the  extraction 
of  tin,  that  the  principal  impurity  of  the  crude  metal  is  iron,  and 
that  to  a  less  extent  metals  such  as  lead,  copper,  tungsten,  etc.> 
may  be  present. 

i.  Liquation.  By  this  process  it  is  chiefly  the  Fe  which  is 
removed  either  in  the  free  state,  or  in  the  form  of  a  rich 
Fe  alloy.  The  impure  tin  is  carefully  heated  so  that  the 
melting-point  of  tin  is  exceeded  as  little  as  possible,  or,  in 
other  words,  so  that  the  tin  alone  is  liquefied,  and  the 
above-mentioned  Sn-Fe  alloy  remains  behind  as  a  solid.  For 
carrying  out  the  process  it  is  customary  to  make  use  of 
small,  simple  reverberatory  furnaces,  the  sloping  hearths 
of  which  are  divided  by  steps  into  two  or  three  sections. 
When  there  are  three  such  sections  the  crude  tin  is  placed 
in  the  middle  one;  the  pure  tin  that  liquates  out  is  allowed 
to  solidify  and  is  then  transferred  into  the  third  section,  which 
is  the  farthest  away  from  the  fire-bridge.  The  tin  that  flows 
away  from  here  is  purified  further  if  necessary.  The  residue 
remaining  in  the  third  compartment,  the  liquation  dross  or 
"  hard-head,"  goes  back  into  the  second  section,  and  all  the 
residues  from  this  in  turn,  including  that  from  the  crude  tin, 
are  placed  in  the  first  section,  which  lies  nearest  to  the  fire. 
The  tin  that  flows  away  from  the  first  compartment,  after 
solidifying,  is  placed  in  the  second,  just  as  that  from  the 
second  goes  to  the  third.  The  dross  from  the  first  com- 
partment is  used  partly  as  an  iron-bearing  flux  in  smelting 


144  METALLURGY 

tin  slags,  and  partly  for  preparing  Pb-Sn  alloys,  by  simply 
melting  it  up  with  Pb.  In  the  smelter  certain  standard 
alloys  are  prepared  and  sold  as  such,  e.g.,  that  with  50% 
Sn  and  50%  Pb)  (common  solder). 

2.  Oxidizing  Fusion.     Although  the  bulk  of  the  Fe,  W  and  a 
part  of  the  Cu  is  removed  by  the  liquation  process,  there 
still  remain  small  amounts  of  these  impurities  in  the  tirj.     They 
may  be  removed  with  the  aid  of  atmospheric  oxygen.     To 
accelerate  the  oxidation  process,  there  is  placed  in  the  iron 
kettle  which  contains  the  molten  tin,  a  pole  of  green  wood 
from  which  vapors  of  water  and  other  decomposition  prod- 
ucts of  the  wood  are  at  once   evolved.     The   melted   metal 
is  stirred  up  by  the  escape  of  these  gases,  so  that  new  portions 
of  it  are  constantly  brought  to  the  surface,  where  they  come 
in  contact  with  the  oxygen  of  the  air.     The  water  vapor  given 
off  by  the  pole  also  tends  to  assist  in  the  oxidation  of  the 
impurities. 

3.  Crystallization  of  Tin.     When  Sn  is  strongly  contaminated 
with  Pb  it  is  almost  impossible  to  effect  a  sharp  and  complete 
separation  of  the  two  metals.     To  be  sure  there  has  been 
no   want   of   attempts   in   this   direction.      Since,   however, 
considerable  quantities  of  Pb-Sn  alloys  are  used  technically, 
it  is  not  at  all  necessary  in  most  cases  to  attempt  such  a 
separation — it  suffices  to  enrich  the  metal.    This,  as  the  experi- 
ments conducted  by  the  author  with  the  aid  of  Mattonet 
have  shown,  can  be  carried  out  in  much  the  same  way  as  in 
the  separation  of  lead  and  silver,  by  the  crystallizing  out 
of  the  lead.     From  the  freezing-point  curve   (Fig.   157)  it 
is  evident  that  the  two  metals  are  soluble  in  one  another 
in    all    proportions    and   that   there   is    a   eutectic    formed 
with  37%   Pb  and  63%  Sn  which  melts  at  about  180°  C. 
If,   therefore,   crude  tin   containing  a  small  percentage   of 
Pb  is  allowed  to   cool  slowly,  it  is  obvious  that  pure  Sn 
will  crystallize  out,  leaving  behind  a  mother  metal  which 
constantly  increases  in  Pb  content,  until   finally   the  com- 
position  of  the  latter    is    that   of    the   eutectic   alloy.     In 


TIN 


145 


practical  work,  the  cooling  will  not  be  carried  out  as  far 
as  this,  but  will  be  stopped  while  the  alloy  is  still  liquid 
and  contains  somewhat  less  Pb  than  the  eutectic  does. 
This  is  drawn  off  in  the  same  way  as  in  the  desilverization 
of  crude  lead,  leaving  behind  an  alloy  containing  a  little  Pb. 
Naturally,  the  process  does  not  yield  perfectly  pure  tin 
the  first  time,  simply  because  the  Sn  crystals  are  more  or 
less  contaminated  with  the  mother  metal,  which  contains  Pb 
and  adheres  to  some  extent  to  the  crystals.  Thus,  to  pre- 


340 

320 
300 
280 
260 
240 
220 


180 


0   10  20  30  I  40  50  60  70  80  90  100$  Pb 
100  90  80  70  |  60  50  40  30  20  , 10   0  jf  Sn 
63$  Sn 

37$  Pb 

FIG.  157. 


pare  perfectly  pure  Sn  it  is  necessary  to  repeat  the  proc- 
ess a  number  of  times,  and,  as  in  the  desilverization  of 
lead  bullion,  there  is  no  difficulty  in  the  work.  The  Sn 
alloy  obtained  by  this  process  of  separation  need  only  be 
brought  to  a  content  of  10%  Pb,  which  is  used  very  extensively 
as  such  for  making  terne-plate. 
Properties  of  Tin: 

SPECIFIC  GRAVITY:    7.3. 

COLOR:   white  with  a  pale  bluish  tinge;   yellow  when  hot. 

MECHANICAL  PROPERTIES:    low  tenacity  and    hardness, 
high  ductility;    most  ductile  at  about   100°;    at   200° 


146  METALLURGY 

so  brittle  that  it  can  be  pulverized.    At  low  temperatures 

it  becomes  changed  into  gray  tin  powder. 
STRUCTURE:  see  Figs.  158,  159,  160. 
MELTING-POINT:  232°  C.  (450°  F.). 
BOILING-POINT:  2100-2200°  C.  (3800-4000°  F.) ;  volatilizes 

perceptibly  at  1200°  C.  (2200°  F.). 
THERMAL   AND   ELECTRICAL   CONDUCTIVITY:    good,   the 

latter  about  0.13  that  of  Ag. 
ALLOYS  with  almost  all  metals.    Many  Sn  alloys  are  valued 


FIG.  158. — Surface  of  Impure  Tin  (X33). 

highly  (with  Cu  =  bronzes;  with  Cu,  Pb,  Sb  =  bearing 
metal,  and  solder;  with  Sb,  Ni,  W,  etc.,  metal  for  fine- 
art  castings).  In  experiments  for  obtaining  alkali  and 
alkaline-earth  metals  by  the  electrolysis  of  their  fused 
salts,  Sn  has  been  found  to  be  a  particularly  good  solvent 
for  these  metals. 

CHEMICAL  BEHAVIOR:  very  stable  toward  O  and  H2O 
at  temperatures  below  its  melting-point;  at  higher 
temperatures  it  oxidizes  readily  and  also  combines 
readily  with  S,  P,  As,  and  Sb.  Cl  and  other  halogens 


TIN 


147 


act  upon  it  even  at  ordinary  temperatures,  forming 
SnCl4,  etc.  Of  the  mineral  acids  HC1  dissolves  the 
metal  best,  H2SO4  only  slowly,  and  HNO3  converts 


FIG.  159. — Deeply  Etched  Under- surf  ace  of  a  Cast  Tin  Ingot. 


FIG.  160. — Large  Grains  caused  by  Tempering  Tin. 

it  into  insoluble  metastannic  acid,  which  is  probably 
a  polymer  of  H2SnO3.  This  dissolves  in  alkali  to  form 
salts  soluble  in  water,  called  stannates.  With  O,  S 
and  on  dissolving  in  acid,  two  series  of  compounds 
are  formed,  those  of  the  stannous  oxide  type  (SnO) 


148  METALLURGY 

called  the  stannous  compounds  (Sn++)  and  those  of 
the  stannic  oxide  type  (SnC^)  called  stannic  compounds 
(Sn+  +  +  +).  In  the  presence  of  strong  bases  Sn 
loses  its  basic  properties  and  unites  with  O  to  form  an 
anion  (stannites  and  stannates).  In  acid  aqueous 
solutions,  as  well  as  in  fusions,  Sn  is  precipitated  by 
Pb  whereas  in  basic  solutions  it  precipitates  Pb. 


ANTIMONY 

Sources 

Natural  Sources : 

NATIVE:   occurs  as  such  rarely. 

As  SULPHIDE:  80283,  which  is  known  as  stibnite;  this  is 

the  most  important  ore  of  antimony. 
As    OXIDE:     80203,    called  senarmontite  or  valentinite. 
Other  Sources : 

Waste  products  from  the  working  up  for  other  metals  of 

ores  containing  antimony,  particularly  in  the  smelting 

of  lead  (dross). 

(A)   Concentration 

Liquation.  The  ready  fusibility  of  Sb2S3  makes  it  possible  to 
melt  out  this  compound  from  its  ores,  leaving  behind,  to  be 
sure,  a  residue  rich  in  Sb,  since  a  considerable  amount  of  Sb-2S3 
(up  to  20%)  entangled  in  the  gangue  remains  behind  in  the 
liquating  apparatus.  In  spite  of  this  difficulty,  such  work  is 
still  carried  out  to  a  considerable  extent  in  Hungarian,  Jap- 
anese, and  Chinese  antimony  districts,  because  the  concentration 
product,  the  so-called  antimonium  crudum,  is  still  much  used 
in  English  antimony  works  for  the  preparation  of  pure  antimony, 
and  in  German  chemical  factories  for  the  manufacture  of 
antimony  preparations.  If  it  were  desired  to-day  to  erect  anti- 
mony works  in  the  vicinity  of  the  mines,  naturally  the  liquation 
process  would  not  be  used  at  all,  because  the  residues  would 
have  to  be  worked  over  considerably  on  account  of  the  large 
amount  of  antimony  they  contain,  and  because  there  is 
not  a  large  amount  of  gangue  to  slag  off  when  the  ores  are 
treated  directly. 

149 


150  METALLURGY 

APPARATUS:   Crucibles  or  Tubes  placed  either  in  air  furnaces  of 
the  simplest  type,  or  upon  the  hearth,  or  in  a  special 
heating  chamber  of  a  reverberatory  furnace.    The  cruci- 
bles are  provided  with  holes  in  the  bottom  and  the  tubes 
are  open  at  both  ends  so  that  the  melted  Sb2Sa  flows  out 
into  collecting  vessels.     Crucibles  heated  in  a  reverber- 
atory furnace  are  placed  over  a  tapping  hole  which  leads 
from  the  bottom  of  the  hearth  to  the  outside.     Reverber- 
atory Furnaces  with  deep  hearths  (see  Fig.  47,  p.  29). 
Oxidizing    Roast    with    Sublimation.       The  ready  volatility 
of  Sb2C>3,  and  the  demand  on  the  part  of  chemical  industries 
for  this  oxide,  has  led  to  an   extensive  practice  of  subjecting 
antimony  ores  to  an  oxidizing  roast,  in  spite  of  the  fact  that 
sulphide  ores  can  be  worked  into  metal  without  any  prelimi- 
nary roasting.     Then  again,  ores  can  be  worked  with  in  this 
manner  from  which  no  antimony  sulphide  could  be  obtained 
by  liquation  and  with  which  the  expense  of  slagging  off  all 
the  gangue  would  be  too  great  on  account  of  the  large  amount 
they  contain.     Now  antimony  on  being  roasted  can  be  con- 
verted into   Sb2O3,  which    has    a    basic    character,    or    into 
Sb2C>5,  which  is  acidic  in  nature.     The  latter  is  formed  more 
readily  in  the  presence  of  basic  substances,  but  as  Sb2O3  is 
itself  basic,  considerable   quantities  of   the  higher  oxide  are 
likely  to  result  unless  special  measures  are  taken  to  prevent 
its  formation  or  to  cause  its  reduction  after  it  has  once  been 
formed.    In  all  cases,  the  aim  is  to  make  Sb2C>3.     Chemical 
industries  demand  this  oxide  in  a  condition  as  free  as  possible 
from  Sb2O5,  because  the  former  is  readily  soluble  in  acid  and 
the  latter  is  not.     But  in  extracting  the  antimony  from  the 
ores  by  the  oxidizing  roast,  it  is  necessary  to  maintain  the  con- 
ditions for  the  formation  of  the  lower  oxide  because  this  is 
readily  volatile,  whereas  Sb2O5  is  not,  and  cannot,  therefore, 
be  removed  as  readily  from  the  gangue  by  volatilization.     If, 
on  the  other  hand,  the  sublimation  product  is  to  be  used  for 
the  manufacture  of  pure  antimony,  it  is    not  necessary  that 
it  should  be  free  from  Sb2O5. 


ANTIMONY  151 

APPARATUS:   Although,  in  metallurgical  literature,  a  number 

of  forms  of  apparatus  and  processes  are  described  which 

are  very  simple,  yet  in  practice  only  low  blast-furnaces 

or  reverberatory  furnaces  are  employed.     In  both  cases, 

it  is,   of  course,   necessary  to  provide  arrangements  for 

properly  catching  and  condensing  the  volatilized  oxide. 

Lixiviation.  It  is  well  known  that  Sb2S3  is  soluble  in  aqueous 

solutions   of  alkali   sulphides  forming  sulphantimonites,  or 

sulphantimonates  upon  the  addition  of  free  sulphur. 

Sb2S3  +  aNa2S  =  2Na3SbS3, 
Sb2S3  +  3Na2S  +  S2  =  2Na3SbS4. 

Chemical  manufacturers  for  a  long  time  have  made  use  of 
these  reactions  in  preparing  artificial,  precipitated  antimony 
sulphide.  Usually  solutions  of  NasSbS4  are  first  prepared  and 
an  intimate  mixture  of  Sb2S3  and  S2,  which  is  regarded  as 
Sb2S5,  is  obtained  on  acidifying.  If,  however,  the  solution  is 
to  be  used  for  the  electrolytic  deposition  of  antimony,  the 
addition  of  S  during  the  lixiviation  is  not  only  unnecessary,  but 
undesirable.  A  solution  of  Na3SbS3  can  be  electrolyzed 
directly,  whereas  one  of  Na3SbS4  must  first  be  treated  with 
an  excess  of  Na2S  or  of  NaOH. 

APPARATUS:  iron  tanks,  conical  at  the  bottom  with  converg- 
ing sides,  and  provided  with  pipes  for  introducing  steam, 
which  serves  to  heat  the  solution  and  to  stir  up  the 
ore  in  it. 

(B)   Extraction 

Reduction  Methods.  For  oxidized  ores  or  metallurigcal  products, 
the  same  methods  are  employed  as  in  the  case  of  the  extraction 
of  Bi  (pp.  in  to  115)  although  in  this  case  it  is  necessary  to 
use  cheaper  fluxes  because  the  price  of  Sb  is  to  that  of  Bi 
as  i :  20. 


152  METALLURGY 

Precipitation   Methods.     This   depends,  as  in  the   like-named 
processes  for  obtaining  Pb  and  Bi,  on  causing  the  reaction 


to  take  place  in  the  melted  mass. 

As  regards  the  apparatus  and  method  of  conducting  the 
operation,  what  was  said  under  Bi  holds  in  the  main 
here  also.  Crucibles  and  reverberatory  furnaces  are 
used.  In  reports  concerning  precipitation  in  rever- 
beratory furnaces  a  process  is  frequently  described  in 
metallurgical  literature  as  the  English  Process,  but  which 
appears  to  be  identical  with  the  practice  that  has  been 
followed  everywhere  that  Sb2Sa  has  been  melted  with 
Fe.  Under  Bismuth  it  was  pointed  out  that  it  is  always 
necessary  in  melting  charges  containing  easily  volatile 
material,  to  provide  a  sufficient  cover  of  fused  material 
so  that  when  a  fresh  charge  is  added  it  will  at  once  sink 
below  the  surface.  In  smelting  rich  antimony  ores,  or  the 
concentration  product  called  antimonium  crudum,  it  is 
clear  that  FeS  will  form  the  principal  constituent  of  the 
slag.  FeS,  however,  is  known  to  be  an  excellent  sol- 
vent for  other  sulphides  and  for  metals,  particularly 
Fe  itself.  Fe,  moreover,  is  the  agent  that  precipitates 
Sb  from  Sb2Sa  in  this  process.  Just  as  in  an  ordinary 
reducing  fusion  it  is  necessary  to  provide  a  reducing 
atmosphere,  and  solid  reducing  agent  in  the  furnace  and 
charge,  so  it  is  clear  that  care  must  be  taken  in  the  pre- 
cipitation process  to  have  enough  Fe  present  in  the  furnace 
before  the  antimony  ore  is  introduced.  Alternately  charg- 
ing with  Fe  and  Sb2Ss  is  prescribed,  therefore,  in  order  that 
the  Fe  may  be  given  time  to  dissolve  in  the  molten  FeS, 
and  thus  attain  a  sufficiently  fine  state  of  division  to  act 
vigorously  enough  upon  the  Sb2S3  with  which  it  sub- 
sequently comes  in  contact. 

Electrolysis.  The  solution  of  Na3SbS3  obtained  by  extraction  with 
alkali  sulphide  (cf.  Concentration  Work,  above)  can  be  electro- 


ANTIMONY  153 

lyzed  directly,  as  has  been  shown  elsewhere,  without  any 
additions  whatever  to  the  bath,  such  as  are  recommended  in  the 
analytical  estimation  of  Sb  by  electrolysis.  Even  although 
the  processes  that  take  place  have  not  been  wholly  explained, 
it  has  been  shown  that  Sb  is  present  in  the  anion  of  the  above 
salt  as  well  as  in  Na3SbS4.  The  separation  of  the  Sb  on  the 
cathode,  therefore,  must  be  the  result  of  a  secondary  reaction; 
first  the  Na  ions  are  discharged  and  then  the  Sb  is  precipitated 
from  the  sulphide.  This  partly  accounts  for  the  fact  that  it  is 
difficult  to  deposit  the  Sb  upon  the  cathode  in  a  dense  condition. 
The  total  result  of  the  electrolysis  may  be  expressed  by  the 
equation, 

2NasSbS3  =  Sb2  (cathode)  +3^282  (anode). 

In  1887  Borchers  established  the  fact  that  large  quantities 
of  Na2S2  are  formed  at  the  anode,  as  well  as  some  NaHS  and 
NasS2O3.  The  last-mentioned  compound  is  formed  readily 
by  the  oxidation  of  Na2S  or  of  NaHS,  and  this  explains  why 
Na2S2O3  can  be  obtained  by  blowing  air  into  solutions  from 
which  antimony  has  been  deposited  electrolytically.  The 
crystallized  salt  is  a  common  commercial  article. 

Special  conditions  for  the  electrolysis  are: 

ELECTROLYZING  VESSELS:  iron  tanks  with  rectangular 
cross-section. 

ELECTROLYTE:    aqueous  solution  of  NaaSbSa. 

ANODES:    lead  plates. 

CATHODES:    iron  plates  and  the  walls  of  the  tanks. 

CURRENT  DENSITY:   10  to  15  amperes  per  sq.ft.  at  the  start; 

E.M.F.,  2  volts;  later  4  to  5  amperes  per  sq.ft. 

By  this  process  not  only  is  the  solvent  recovered  in  a  com- 
mercial form,  but  also  the  sulphur  of  the  ore. 

Inasmuch  as  it  is  difficult  to  prepare  antimony  solutions 
perfectly  free  from  Fe,  there  is  always  more  or  less  FeS  formed 
on  the  walls  of  the  electrolyzing  tank,  and  upon  the  surface  of 
the  cathode,  and  this  falls  to  the  bottom  with  the  metal  that 


154 


METALLURGY 


drops  off;   hence  this  process  does  not  yield  directly  a  perfectly 
pure  metal,  but  rather  one  that  contains  Fe. 

(C)   Antimonial  Lead,  Hard  Lead 

In  the  purification  of  lead  bullion  (cf.  Silver,  Desilverization, 
and  Lead)  by  an  oxidizing  fusion,  one  of  the  first  oxidation  prod- 
ucts withdrawn  is  usually  a  mixture  of  PbO  and  Pb3(SbO4)2. 
We  have  already  seen  in  the  above  cited  sections  that  this  product 
is  next  subjected  to  a  liquation  during  which  a  reducing  flame 


<J. 

600 
500 

400 

O 
300 
R 

631 

^ 

-"" 

^ 

•^ 

326° 

^^ 

^ 

^ 

X 

NP^ 

*^~ 

247° 

*  —  13- 

Q 

10    20    30    40    50    60    70    80    90   1(X 

FIG.  161. 

is  maintained  and  small  quantities  of  charcoal  powder  are  spread 
over  the  charge.  By  this  treatment  it  is  aimed  to  cause  the  separa- 
tion of  mechanically  enclosed  Pb  granules;  but  by  means  of  the 
reducing  flame  and  the  charcoal  powder  a  part  of  the  PbO  is 
reduced  at  the  same  time  and  the  residue  enriched  in  Sb.  The 
further  working  up  of  the  antimony  skimmings  is  carried  out,  as 
was  explained,  by  a 

Reducing  Fusion  in  small  shaft  furnaces,  such  as  serve  for 
the  smelting  of  lead  ores.  In  smelters  in  which  but  little 
of  this  by-product  is  obtained,  low  blast  furnaces  are  also 
used  for  working  up  the  litharge  when  convenient. 

As  a  reduction  product  an  Sb-Pb  alloy  is  obtained  with  14 
to  20%  Sb.  The  old  assumption  that  hard  lead  is  a  chemical 


ANTIMONY  155 

compound,  Pb3Sb2,  is  untrue.  Pb  and  Sb,  as  is  evident 
from  Fig.  161,  are  only  soluble  in  one  another  as  solids,  and 
form  a  eutectic  containing  13%  Sb,  which  melts  at  247°  C. 

(D)   Antimony  Refining 

The  chief  impurities  of  crude  antimony  are  Fe  (7  to  10%)  S 
(up  to  i%)  and,  as  a  rule,  only  small  amounts  of  other  metals. 
The  refining  of  the  metal,  therefore,  consists  mainly  in  the  removal 
of  Fe,  but  during  the  process  the  other  impurities  disappear  for 
the  most  part.  Pb  is  very  difficult  to  remove  from  Sb;  ores 
showing  a  high  percentage  of  Pb  must  be  converted  by  separate 
smelting  operations  into  an  Sb-Pb  alloy  (hard  lead)  and  crude 
Sb.  The  removal  of  the  Fe  and  S  is  effected  by  the  following 
Operations : 

1.  Sulphurizing  Fusion  of  the  crude  metal  with  Sb2Ss    (8  to 
12%)   and   a   little   common   salt,   NaCl    (4   to   5%)    in   a 
crucible  or  reverberatory  furnace  of  the  same  construction 
as  described  under  Bismuth   (cf.  Figs.    130  to  137,  pp.   112 
and  114). 

By  means  of  this  fusion  the  metal,  is  brought  to  a  purity 
of  98  to  99%  Sb  with  a  low  content  of  Fe  and  S.  These 
last  impurities  are  removed  by  the  so-called 

2.  Starring.     The  metal,  together  with  a  strongly  basic  flux, 
composed   of  about   3   parts   soda   ash.   or  potash,   and   2 
parts  Sb2Ss,  is  again  melted  in  either  a  crucible  or  rever- 
beratory furnace.     In  the  above  case  the  Sb2$3  was  used  as 
an  agent  for  the  removal  of  Fe,  but  here,  in  the  presence  of 
alkali,  it  also  had  a  tendency  to  form  sulphantimonate  and 
acts,  therefore,  as  a  desulphurizing  agent  as  well.     During 
this  smelting  it  is  of  course  necessary  to  take  particular  pains 
to  avoid  having  the  metal  come  in  contact  with  iron;    the 
iron  instruments  used  for  stirring  and  for  ladling  are,  for 
this  reason,  always  given  a  coat  of  enamel  or  whitewash,  as 
far  as  there  is  any  danger  of  their  coming  in  contact  with  the 
metal.     If,  in  the  ladling  or  pouring  out  of  the  purified  metal 


156  METALLURGY 

from  the  vessel  into  the  whitewashed  iron  molds,  care  is  taken 
to  scoop  up  and  pour  out  enough  slag  to  form  a  protective 
cover,  then  the  crystallization  of  the  Sb,  which  is  associated  with 
the  cooling  of  the  molten  metal,  will  start  at  the  walls  of  the 
mold.  The  crystals  grow  beneath  the  badly  conducting  slag 
layer  and  the  last  portions  of  the  liquid  collect  at  the  sur- 
face, because  they  are  there  most  protected  from  cooling; 
this  liquid  will  be  constantly  drawn  back  by  the  crystalline 
framework  which  diminishes  in  surface  area  during  the  further 
cooling.  Thereby  the  crystal  dendrites  that  first  form  in 
it  are  pushed  forcibly  through  the  surface  of  the  melt  and 
"  stars  "  are  formed  there,  which  are  regarded  by  buyers  as 
a  sign  of  purity.  This  phenomenon  has  given  the  name  of 
"starring"  to  the  entire  operation. 

Electrolytic  Production  of  Pure  Antimony.  This  has  been 
attempted  under  different  working  conditions,  but  up  to 
the  present  time  such  processes  have  not  been  adopted  on 
a  large  scale.  Among  such  experiments  may  be  mentioned 
the 

Electrolysis  of  Crude  Antimony  with  SbCl3.KCl  solution  as 
electrolyte  (Sanderson)  for  working  up  auriferous  antimony. 
Electrolysis  of  an  SbCl3.FeCl2  Solution,  using  insoluble  anodes, 
with  the  idea  that  FeCl2  will  be  converted  into  FeCl3  at  the 
anode,  and  that  the  latter  will  serve  as  solvent  for  the 
lixiviation  of  Sb2S3  ores. 

ELECTROLYSIS:    6  FeCl2  +  2SbCl3  =  Sb2  +  6FeCl3. 
LIXIVTATION:   6  FeCl3  +  Sb2S3  =  6FeCl2  +  2SbCl3  +  3S: 

The  lixiviation,  as  well  as  the  electrolysis,  of  such  chloride 
solutions    involve    difficulties    and   high    cost    in   providing 
sufficiently  durable  apparatus  and  in  replacing  the  cathode 
metal  upon  which  the  FeCl3  has  a  solvent  action. 
Properties  of  Antimony: 

SPECIFIC  GRAVITY:  6.7. 

MECHANICAL  PROPERTIES:  brittle. 

COLOR:   white  with  high  metallic  luster. 

STRUCTURE:    pure  antimony  shows  a  large-leafed,  crys- 


ANTIMONY 


157 


talline  structure,  the  surface  of  the  so-called  "  antimony 
button "  showing  star-shaped  outlines  (cf.  Figs.  162, 
163,  164). 


FIG.  162. — Surface  (Xi5). 


FIG.  163. — Internal  Grain  Structure. 

MELTING-POINT:    631°  C.  (1168°  F.),  readily  volatile  at 
higher  temperatures. 


158  METALLURGY 

THERMAL  AND  ELECTRICAL  CONDUCTIVITY:  slight,  the 
latter  about  0.035  that  °f  Ag.  In  the  thermo-electric 
potential  series,  it  is  the  most  negative  metal  used  in  the 
construction  of  thermo  elements. 

ALLOYS:  particularly  with  the  metals  Pb,  Bi,  Sri,  Cu,  Ni, 
Fe,  Cu;  with  the  last  lour  metals  chemical  compounds 
are  also  formed. 

CHEMICAL  BEHAVIOR.  At  ordinary  temperatures  it  resists 
well  the  action  of  O  and  H-O.  Above  its  melting-point 
it  oxidizes  readily  and  burns  to  Sb2O3  and  Sb2O5, 
or  to  Sb2C>4,  which  may  be  regarded  as  containing  half 


FIG.  164. — Internal  Dendritic  Structure  (X4o). 

a  molecule  of  each  of  the  other  oxides.  Sb2O3  shows 
both  acid  and  basic  properties,  but  Sb2Os  has  an  acid 
character.  Sb  combines  directly  with  Cl  to  form  SbCl5 
or  SbCl3.  HC1  dissolves  Sb  slowly.  H2SO4  oxidizes 
Sb  to  Sb2O3  and  an  excess  of  this  acid  forms  a  sulphate, 
which  hydrolyzes  readily.  HNO3  oxidizes  Sb  to  Sb2O5- 
Sb  shows  a  great  affinity  for  S ;  the  sulphides  Sb2S3  and 


ANTIMONY  159 

Sb2Ss  (the  latter  is  unstable  in  the  free  state)  have  an 
acid  character  and  unite  with  alkali-  and  alkaline-earth 
sulphides  to  form  sulphantimonites  and  sulphantimo- 
nates,  which  are  readily  soluble  in  water.  Tn  the  fused 
state  the  affinity  for  S  is  so  great  in  the  presence  of 
alkali  monosulphides  that  the  sulphantimonites,  e.g., 
Na^SbSs,  will  withdraw  S  from  even  Sb2S3,  with  sepa- 
ration of  Sb,  to  form  Na3SbS<i,  which  represents  the 
pentavalent  condition  of  antimony. 


NICKEL 

Sources 

Natural  Sources : 

Native  in  meteoric  iron,  up  to  20%. 
SULPHIDE  ORES. 

MILLERITE:   NiS  (rare). 

PENTLANDITE:   (FeNi)StoNiS.2FeS. 

HORBACHITE:   (FeNi)2S3. 

LINN^EITE:   (NiCo)3S4. 

GERSDORFFITE:  NiAsS. 


usually 
with  other 
sulphides, 
pyrite  and 
pyrrhotite. 


ULLMANNITE:   NiSbS. 
ARSENIDES  AND  ANTIMONIDES. 

NICCOLITE:  NiAs. 

CHLOANTHITE:   NiAs2. 

BREITHAUPTHITE  :  NiSb. 
SALTS. 

MORENOSITE:    NiSO4  +  7H2O. 

ANNABERGITE  :    Ni3 ( AsO4) 2  f  8H2O. 

ZARATITE:  a  basic  carbonate. 

GARNIERITE:  a  hydrosilicate  of  nickel  and  magnesium. 
Other  Sources: 

Nickeliferous  rocks,  speisses  and  slags. 
Metal  clippings,  metal  waste,  old  metals. 

(A)   Concentration  Methods 

Nickel  in  its  ores  is  often  associated  with  large  quantities  of 
gangue  in  such  variety  of  ways  that  a  direct  smelting  of  the  ore, 
wherever  smelting  comes  into  consideration  at  all,  is  out  of  the 
question. 

160 


NICKEL  161 

As  in  the  case  of  copper  ores,  and  according  to  the  same  prin- 
ciples (cf.  pp.  60  to  90),  it  is  possible  to  carry  out  a  concentrating 
process  consisting  of  roasting  and  smelting  operations,  but  in 
some  cases  it  is  more  advantageous  to  employ  a  concentrating 
procedure  which  consists  of  roasting,  lixiviation  and  precipita- 
tion. 

Of  the  numerous  proposals  for  concentrating  nickel  ores, 
the  choice  is  determined,  in  the  first  place,  by  the  presence  or 
absence  of  copper,  or  by  whether  it  is  desired  to  effect  a  separa- 
tion of  nickel  from  copper;  and,  in  the  second  place,  by  the  ques- 
tion whether  it  is  worth  while  to  attempt  a  separation  from  cobalt. 
The  removal  of  the  other  metals  which  are  likely  to  be  present 
in  nickel  ores  offers  no  special  difficulties  and  in  no  way  influences 
the  work  of  concentrating  the  nickel. 

Concentration  without  Separation  of  Copper.  By  this  work 
all  copper-free  ores  can  be  handled,  but  if  copper  is  present 
it  is  useful  only  when  it  is  desired  to  prepare  by  smelting  an 
alloy  of  copper  and  nickel  rather  than  pure  nickel.  Next 
to  nickel,  copper  has  the  greatest  affinity  for  sulphur.  For 
this  reason  the  nickel  ore  can  be  concentrated  by  roasting 
and  smelting  to  a  matte  just  as  was  described  under  Copper. 
In  the  case  of  nickel,  however,  there  is  another  possibility  for 
concentrating,  based  upon  the  fact  that  nickel  has  the  greatest 
affinity  for  arsenic  of  all  the  different  metals  that  come  into 
consideration  here.  Of  course  this  method  of  smelting,  the 
so-called  speiss  smelting  (the  arsenides  of  Ni,  Co  and  Fe  form 
what  is  called  speiss},  is  applicable  only  when  the  ores,  or 
metallurgical  products  at  hand  contain  arsenic;  in  the  case 
of  arsenic-free  ores  the  aim  must  always  be  to  form  a  matte. 

The  mode   of   operating    and    the    apparatus   are    almost 
exactly  the  same  as  in  the  case  of  copper  concentration,  namely 
i.  Oxidizing    Roast,    employed    with    sulphide   ores    for   the 
same   reasons    as    in    the    case    of    sulphide    copper    ores, 
and  with    arsenide    ores  in    order    to    reduce    the  As-con- 
tent as  far  as   i   As   to   2    Ni,    but   not   any   farther.      A 
further   reduction    in    the    amount    of    As    would    lead    to 


162  METALLURGY 

Ni  losses  in  the  slag  during  the  succeeding  work;  in  smelt- 
ing for  crude  matte  or  crude  speiss,  however,  the  metal  con- 
tent of  the  resulting  slag  must  be  sufficiently  low  to  permit 
discarding  of  the  slag. 

Apparatus  for  carrying  out  the  roast,  cf.,  Copper. 
PRODUCTS:    SO2    (eventually   to  be   worked  into   H2SO4); 

As2O3  to  be  worked  into  white  arsenic  by  condensation 

and  resublimation ;   roasted  ore. 

2.  Formation  of  First  Matte  or  First  Speiss.     The  roasted  ore 
from  the  above  treatment,  or  oxidized  ore  with  the  addition 
of  sulphide    (pyrites,  tank  waste,   sulphates)   together  with 
rich  Ni  slags  from  the  following  operations  and  other  waste 
products,  are  placed  in  a  shaft  or  reverberatory  furnace  and 
subjected  to  a  reducing  smelt. 

PRODUCTS:  a  slag  low  in  metal  content  which  can  be  dis- 
carded unless  there  is  demand  for  it  as  building  material, 
etc.  A  crude  matte  containing  up  to  40%  Ni  (perhaps 
Ni  and  Cu)  and  when  low  in  Ni,  up  to  40%  Fe.  In  the 
case  of  arsenical  raw  materials,  a  speiss  is  obtained  with 
from  40  to  50%  Ni. 

3.  Oxidizing  Roast  of  the  Crude  Matte,  or  Crude  Speiss:     the 
purpose  of  this  operation  is  the  same  as  under  i. 

APPARATUS:  reverberatory  furnace. 

PRODUCTS:  SO2,  As2O3  as  under  i,  roasted  matte  or  speiss. 

4.  Concentration  of  the  Matte  or  Speiss  is   carried    out  with 
acid  fluxes  in  shaft  or  reverberatory  furnaces  as  in  the  copper 
industry. 

PRODUCTS:  nickeliferous  slag  which  must  be  again  added 
to  the  charge  in  the  smelting  of  crude  matte,  and  a  matte 
richer  in  Ni  which  still  contains  considerable  Fe;  this 
must  be  subjected  to 

5.  Blowing  in  Converters,  by  which  means  a  product  containing 
76%  Ni  is  obtained. 

The  speiss  from  concentration  is  not  refined,  as  a  rule, 
but  goes  to  the  metal  extraction  processes  with  a  Ni  con- 
tent of  65  to  70%  Ni. 


NICKEL  163 

As  regards  the  constitution  of  nickel  matte,  most  metallurgical 
text-books  give  incorrect  statements.  The  true  constitution 
was  established  by  Bornemann  in  1907.  The  sulphide  NiS  is  not 
capable  of  existence  under  the  conditions  prevailing  in  the  smelting 
of  matte  and  is  never  to  be  found  even  in  the  relatively  rich 
mattes.  According  to  the  concentration  of  Ni,  Fe,  and  S  in  the 
matte,  there  may  be  present  combinations  of  Ni2S  with  FeS, 
of  Ni3S2  with  FeS,  free  Ni3S2,  and  when  still  less  S  is  present 
there  may  be  present  a  solution  of  Ni  in  NisS2. 

Bornemann  obtained  the  following  combinations:  (FeS)2Ni2S, 
which  crystallizes  from  molten  matte  at  686°  C.  but  is  stable  only 
at  high  temperatures  and  changes  at  575°  into  (FeS)3(Ni2S)2 
with  loss  of  FeS.  At  lower  temperatures  FeS  is  again  taken  up 
to  form  the  compound  (FeS)4Ni2S. 

The  poorer  the  matte  is  in  FeS,  the  more  of  the  compound 
Ni3S2  appears,  as  far  as  the  concentration  of  the  sulphur  permits. 
In  mattes  that  are  free  from  FeS  there  are  found,  according  to  the 
concentration  of  the  sulphur,  Ni3S2;  Ni2S,  and  free  Ni,  the  last 
two  being  held  in  solution  by  Ni3S2.  Every  experiment,  in  which 
it  was  attempted  to  form  the  compound  NiS,  so  often  mentioned 
as  a  constituent  of  nickel  matte,  was  unsuccessful.  Under  the 
conditions  prevailing  in  matte  formation,  the  compound  Ni3S2 
represents  the  most  sulphur  that  can  be  made  to  combine  with 
nickel. 

Furthermore,  most  statements  in  metallurgical  literature  pre- 
vious to  1907  concerning  the  constitution  of  nickel  speiss  are 
likewise  erroneous.  Friedrich  has  been  able  to  identify  with 
certainty  only  two  compounds  between  Ni  and  As,  namely,  Ni5As2 
and  NiAs;  he  believes,  however,  that  the  existence  of  a  compound 
Ni3As2  is  probable.  The  latter  is  formed,  judging  from  the  behav- 
ior of  the  melts  examined,  from  mixed  crystals  of  Ni5As2  with 
As,  and  the  compound  NiAs  is  often  formed  only  after  the  com- 
plete solidification,  under  strong  supercooling,  of  a  eutectic  cor- 
responding to  this  composition.  By  inoculation  (throwing  in  a 
crystal)  it  is  possible  to  prevent  the  supercooling  and  bring  about 
the  solidification  of  the  eutectic  and  the  carrying  out  of  the  reaction 


164  METALLURGY 

according  to  which  Ni3As2  is  formed  at  a  temperature  lying  above 

the  eutectic  point. 

Concentration  of  Nickel  and  Elimination  of  Copper  by  the 
Formation  of  Matte  and  Speiss.  If  ores,  metallurgical 
products,  or  metallic  scraps  are  at  hand  which  contain  Ni,  Fe,  Cu, 
S,  and  As,  a  separation  of  the  Ni  and  Cu  can  be  effected,  although 
not  very  sharply,  if  the  roasting  and  smelting  work  is  conducted 
in  such  a  way  that  the  material  as  it  comes  to  the  smelter  con- 
tains enough  sulphur  to  form  a  copper  matte  and  enough  As 
to  form  a  nickel  speiss.  On  account  of  the  great  affinity  of 
Cu  for  S  and  of  Ni  for  As,  the  greater  part  of  the  Cu  will*  then 
be  in  the  matte  and  the  greater  part  of  the  Ni  will  be  in  the 
speiss.  Although  the  matte  and  the  speiss  dissolve  in  one 
another  to  a  slight  extent,  still  for  the  most  part  they  are  obtained 
from  the  smelting  in  a  separate  state.  The  copper  matte  is, 
in  other  words,  not  free  from  Ni  and  the  Ni  speiss  is  not  free 
from  Cu,  so  that  the  work  must  be  repeated  with  both  of  these 
products.  The  methods  used  in  carrying  out  this  work  do  not 
differ  essentially  from  those  described  above  for  producing 
matte  and  speiss. 

Concentration  of  Nickel  and  Elimination  of  Copper  by  Smelt- 
ing to  a  Matte.  This  is  employed  in  working  up  sulphide 
ores  carrying  Ni  and  Cu  (pyrrhotite)  of  which  large  deposits 
occur  in  the  Sudbury  district,  Ontario,  Canada,  and  in  South- 
ern Norway.  This  ore  also  occurs  in  Southern  Germany,  but 
the  deposits  there  have  never  been  worked  to  any  extent.  For 
the  last  forty  years  a  process  has  been  used  in  some  smelters 
which  is  quite  complicated  and  consists  of  the  following  opera- 
tions : 

1.  Oxidizing  Roast  of  the  Ore.  1 

\  as  under  i  and  2,  pp.  161-2. 

2.  Smelting  for  First  Matte. 

3.  Matte   Smelting  with  the  addition    of   Na2SO4   and   char- 
coal. 

APPARATUS:  shaft  furnace  usually  constructed  with  water 
jackets,  cf.  Johnson  furnace,  p.  80,  Figs.  105  and 
106. 


NICKEL  165 

PRODUCTS:  Besides  a  waste  slag,  there  is  a  Cu-Fe-Na 
matte  and  a  specifically  heavier  Ni-Fe  matte.  After  being 
tapped  into  slag  pots,  the  two  mattes  separate  into  layers. 
The  Cu-Fe-Na  matte  after  solidifying  forms  the  so-called 
"  tops,"  and  can  be  broken  away  from  the  other  matte 
which  constitutes  the  "  bottoms."  The  tops  resulting 
from  this  treatment  are  not  yet  poor  enough  in  Ni  to  be 
worked  for  Cu,  and  likewise  the  bottoms  contain  too  much 
Cu  to  be  worked  directly  for  Ni.  It  is  necessary,  there- 
fore, to  treat  both  of  these  products  again. 

4.  Smelting    of   the    Tops.       The   copper  matte   which   con- 
tains  sodium   is   first   allowed    to    weather    somewhat,  the 
desired  action  being  accelerated  by  wetting  with  the  leach- 
ings  from  No.  7  (see  below).     Then,  with  the  addition  of 
some  raw  matte  (from  2)  it  is  subjected  to  a  reducing  smelt 
in  a  shaft  furnace. 

PRODUCTS:  waste  slag;  tops  consisting  of  a  matte  rich 
in  Cu  but  poor  in  Ni,  and  a  specifically  heavier  matte 
which  has  a  Ni  content  corresponding  to  about  that  of 
the  former  bottoms  with  which  it  is  smelted  again. 

5.  Smelting   of    the    Bottoms.      The  heavy  mattes,  obtained 
by  operations  3  and  4,  are  smelted   with  sodium  sulphate 
and  charcoal  in  exactly  the  same  way  as  in  the  third  opera- 
tion. 

PRODUCTS:  Besides  a  waste  "slag,  a  specifically  light  matte 
which  corresponds  in  Cu  content  to  the  tops  of  opera- 
tion 3,  as  well  as  to  those  of  operation  4;  with  both  of 
these  it  is  again  smelted.  The  specifically  heavier  matte 
forms  a  concentrated  bottom  which  is  rich  in  Ni  and 
poor  in  Cu. 

6.  The  concentrated  tops  are  leached  with  water. 
PRODUCTS:   A   solution   of  sodium   sulphide   and   oxidation 

products;  on  evaporating  to  dryness,  a  mixture  of  Na2S 
and  other  Na  salts  is  obtained  which  can  be  used  again 
as  a  flux  in  the  third  operation.  The  leached  residue 
consists  chiefly  of  Cu2$  together  with  practically  all  the 


166  METALLURGY 

Au  and  Ag  originally  present;  it  is  worked  tip  in  the  usual 
way  for  Cu  and  the  precious  metals. 

7.  The  concentrated  bottoms  are  subjected  to  a  chloridizing 
roast  carried  out  in  such  a  way  that  it  yields,  after  treatment 
with  water,  the  following 

PRODUCTS  :  A  solution  containing  the  chlorides  of  the  platinum 
metals  and  copper;  these  metals  are  recovered  by  precipi- 
tation in  the  usual  manner.     The  leached  residue  contains 
a  mixture  of  NiO  with  a  little  SiC>2,  S,  Cu,  Fe  and  Pt.    For 
the  method  for  working  this  up,  see  Extraction. 
Concentration  by  Leaching  and  Precipitation,   if  necessary, 
with  the  aid  of  roasting.     Directly  applicable  to  ores,  inter- 
mediate   products  and  by-products  which  contain  the  Ni  in 
the  form  of  easily  soluble  salts,  or  free  oxides  and  sulphides. 
After   a   preliminary   roasting   it   is   particularly   suitable    for 
sulphide  and  arsenide  ores  and  intermediate  products  which 
also  contain  besides  nickel  other  metals  worth  recovering  (Cu> 
Co  and  precious  metals).     Contrary  to  the  many  statements  in 
metallurgical  literature,  it  is  to  be  emphasized  that  silicates 
are  absolutely  unsuited  for  such  work. 

An  Old  Process  of  Roasting,  Lixiviation,  and  Precipitation  was 
formerly  used  universally  for  working  up  mattes  and  speiss 
which  contained  Co,  Cu  and  other  metals  besides  Ni.  It 
consists  of  the  following  separate  operations: 

1.  DEAD  ROASTING  to  Oxides. 

2.  SOLUTION  OF  THE  ROASTED  RESIDUE  in  acids  (HC1  or 
H2S04). 

3.  PRECIPITATION  by  H2S  or  by  a  soluble  sulphide  (Na2S, 
etc.). 

PRODUCTS:  Sulphides  of  Cu,  Pb,  Bi,  Sb,  As.  These  sul- 
phides are  utilized  in  the  works  according  to  the  predomi- 
nant metal. 

4.  OXIDATION  of  the  residual  solution  with  chloride  of  lime 
and  neutralization  with  milk  of  lime.     First  of  all,  the 
iron  is  precipitated  as  Fe(OH)3.     After  the  greater  part 
of  the  iron  has  come  down,  the  solution  is  filtered  and  the 


NICKEL  167 

addition  of  chloride  of  lime  and  milk  of  lime  continued. 

The  cobalt  is  thus  precipitated  as  hydrated  Co2O3. 
5.  PRECIPITATION  with  Milk  of  Lime  or  Soda. 

PRODUCT:  Ni(OH)2  or  NiCO3. 

Herrenschmidt's  Process  for  working  up  ores  carrying  Co  and 
Ni.  This  process  is  used  on  Herrenschmidt's  own  property 
for  asbolite  (18%  Mn2O3,  3%  CoO,  1.25%  NiO,  30%  Fe2O3, 
8%  SiO2,  5%  A12O3,  i%  CaO,  i%  MgO).  It  consists  of 
the  following  operations: 

1.  OXIDIZING  ROAST. 

2.  TREATMENT  of  the  roasted  residue  with  ferrous  sulphate 
solution  and  introduction  of  steam  and  air. 
PRODUCTS:  CoSO4  and  NiSO4  in  solution. 

3.  PRECIPITATION  of  the  solution,  which  is  kept  weakly  acid, 
with  Na2S. 

PRODUCTS  :  a  small  amount  of  CoS  and  NiS. 

4.  THE  SULPHIDES  are  filtered  off,  dried  and  given  a  sul- 
phatizing  roast. 

5.  THE  ROASTED  PRODUCT  is  dissolved  in  water,  and  after 
treatment  with  CaCl2  solution,  CoCl2  and  NiCl2  remain  in 
solution  while  CaSO4  is  precipitated. 

6.  THE  SOLUTION  obtained  in  5  is  divided  according  to  the 
ratio  of  Co  to  Ni.     One  part  of  this  solution  is  completely 
precipitated    with    Ca(OH)2    solution.      The    precipitate 
is  converted  into  Co2O3  and  Ni2O3  by  means  of  air  and 
chlorine.     This  precipitate  is  mixed  with  the  remainder 
of  the  original  solution  so  that  the  Ni2O3  reacts  with  the 
CoCl2  in  solution  to  form  Co2O3  and  soluble  NiCl2. 

7.  FILTRATION:  the  filtered  precipitate  when  dried  is  Co2O3 
which  can  be  sold  as  such. 

8.  THE  FILTRATE,  containing  NiCl2  in  solution,  is  precipitated 
by  Ca(OH)2. 

9.  FILTRATION:  the  filtered  precipitate  of  Ni(OH)2   is  dried 
and    ignited,    whereby    it    is    converted  into  NiO.    This 
is  worked  up  as  described  below  under  B.     The  liquors 
containing  CaCl2  can  be  used  again  in  Operation  5. 


168  METALLURGY 

Borchers  and  Warlimont's  Process  accomplishes  in  a  simple 
manner  a  working  up  of  Ni  ores  that  carry  Co  and  Cu. 
It  consists  of  the  following  operations: 

1.  SULPHATIZING  ROAST  at  a  temperature  of  about  500°  C, 
in  rotating  iron  drums,  into  which   the  air  necessary  for 
roasting  is  led.     The  drums  are  heated  externally. 

PRODUCTS:  SO2,  which  can  be  utilized  in  the  usual  way, 
and  a  roasted  residue  in  which  the  Co  and  Cu  are  present 
almost  entirely  as  sulphates,  the  Ni  partly  as  Ni3S2, 
partly  as  NiO  and  only  to  a  slight  extent  as  NiSO4, 
while  the  Fe  is  converted  for  the  most  part  into  Fe2Os, 
although  some  FeSO4  and  Fe(SO4)s  remain. 

2.  EXPOSURE  of  the  roasted  product  to  the  air  for  several 
days,  meanwhile  working  it  about  and  wetting  it.    Hereby 
the  FeSO4  that  is  present  converts  any  unchanged  Cu2S 
into  CuSO4. 

3.  LEACHING  with  slightly  acid  water. 

PRODUCTS:  a  solution  of  CuSO4  and  CoSO4,  which  is  only 
slightly  contaminated  with  NiSO4  and  FeSO4,  and  a 
residue  which  contains  most  of  the  Ni  together  with  the 
original  gangue  of  the  ore  and  the  Fe2O3  that  has  been 
formed. 

4.  PRECIPITATION  with  iron. 

PRODUCTS:  cement  copper,  concerning  the  working  up  of 
which  see  Copper,  and  a  weakly  acid  solution  containing 
all  the  Co  of  the  ore  and  a  small  amount  of  Ni,  both  in 
the  form  of  sulphates. 

5.  PRECIPITATION  of  the  solution  obtained  in  4,  using  CaS 
as  precipitant;    the  latter  is  a  by-product  in  the  smelting 
of  concentrated  matte  with  CaO  -I-  C. 

PRODUCTS:  a  mixture  of  CoS  with  a  little  Ni3S2  which  is 
dried  and  used  in  the  cobalt  industry.  The  liquor  obtained 
here  may  be  evaporated  to  dryness  and  the  residue  used 
as  a  flux  in  operation  i. 

6.  Matting  of  the  leached  residue  obtained  in  Operation  3 
adding  to  the   charge,  when    necessary,  substances    con- 


NICKEL  169 

taining  S  and  free  from  Cu.  At  the  same  time  other 
nickeliferous  slags  and  by-products  obtained  in  subse- 
quent operations  may  be  added  to  the  charge. 
PRODUCTS:  waste  slag  and  first  nickel  matte  which  is 
worked  up  further  as  described  under  A  on  page  162.  As 
regards  the  method  of  obtaining  metallic  nickel  see  the 
following  descriptions  of  Extraction  and  Refining. 

(B)    Extraction 

The  choice  of  a  process  for  working  up  the  products  obtained 

by  concentration  work  is  determined  by  the  character  of  the 

material   (whether  a  concentrated  matte,  a  concentrated   speiss, 

or  an  oxide),  the  character  of  the  finished  product  (whether  pure 

nickel,  or  a  nickel  alloy),  and  finally,  by  the  availability  of  a  cheap 

fuel  or  other  source  of  energy  (water-power). 

The  Roast-Reduction  Process  has  in  the  past  been  most  used 

because    the    concentration    work,   as    described    above,   has 

yielded  chiefly  a  concentrated  matte  or  speiss.     This  process 

comprises  the  following  operations: 

i.  Dead  Roast.  Matte  or  speiss  is  roasted  with  great  care 
in  a  reverberatory  furnace  until  the  last  traces  of  S  or 
As  have  been  oxidized  and  removed.  Sometimes  it  is 
necessary  to  add  a  little  Na2CO3  or  NaNOs  (or  a  mixture 
of  both  salts)  in  order  to  remove  the  last  traces  of  As. 
APPARATUS:  reverberatory  furnaces. 

PRODUCTS:  SO->,  As2C>3,  which,  however,  on  account  of  the 
low  concentration  in  which  they  are  present  in  the  gases, 
cannot  be  utilized;  hence,  wherever  this  method  is  employed 
its  effect  is  very  injurious  to  vegetation  in  the  adjacent 
country.  The  roasted  residue  consists  almost  entirely 
of  NiO.  In  case  one  or  both  of  the  above-mentioned 
Na  salts  were  used  to  effect  the  removal  of  the  As,  the 
residue  also  contains  Na3AsO4,  which  can  be  extracted 
with  water. 
2.  Reduction.  For  this  work,  there  is  the  residue  obtained 


170  METALLURGY 

under  i  as  well  as  oxidized  products  obtained  in  concen- 
trating work  as  described  under  A.  The  reduction  may 
consist  either  of  a  reducing  roast  or  of  a  reducing  smelt. 
The  reduction  by  means  of  a  reducing  roast  is  effected 
by  taking  the  NiO  obtained  in  previous  operations,  stirring 
it  up  with  flour  paste  to  a  plastic  mass,  rolling  it  into  plates, 
and  cutting  these  lengthwise  and  crosswise  into  small 
cube-shaped  briquettes,  which,  after  being  dried  in  retorts 
or  crucibles,  are  embedded  in  charcoal  powder  and  heated 
within  the  same  vessels  to  the  reduction  temperature  of 
NiO,  usually  using  producer  gas  as  fuel.  It  is  not  intended 
hereby  to  effect  a  complete  fusion,  but  merely  to  weld 
together  the  reduced  Ni;  the  cubes,  naturally,  are 
smaller  in  size  than  the  cubes  of  NiO  that  were  intro- 
duced into  the  reduction  furnace.  The  "  cube  nickel  " 
thus  obtained  is  preferred  in  this  form  by  buyers  for  many 
purposes.  It  is,  for  example,  suitable  for  purposes  of 
alloying,  but  it  is  not  suitable  for  working  into  nickel  wares 
on  account  of  the  presence  of  carbides  as  well  as  some 
NiO,  which  impurities  cause  brittleness.  A  direct  reduc- 
ing smelt  of  nickelous  oxide,  which  may  take  place  in 
a  crucible  furnace  or  in  a  shaft  furnace,  is  usually 
employed  when  the  oxidized  concentration  product  still 
contains  metals  that  should  either  be  eliminated  by  fur- 
ther separation  processes  (cf.  Nickel  Refining),  or  when  it 
contains  large  amounts  of  other  oxides  (CuO),  which  it 
is  desired  to  reduce  at  the  same  time  so  that  a  nickel  alloy 
will  be  formed. 

Roast-Reaction  Smelting  in  reverberatory  furnaces  or  in 
converters  does  not  give  rise  to  metal,  as  is  the  case  when 
working  with  copper  products.  The  reason  for  this  was 
found,  in  the  author's  laboratory,  to  be  as  follows:  The 
reaction 

Ni3S2  +  4NiO  =  ;Ni  +  2SO2 
takes  place,  to  be  sure,  at  temperatures  above  1400°  C,  but 


NICKEL  171 

only  very  slowly  even  when  the  heat  is  raised  above  1600°  C, 

and  with  high  oxygen  concentration  in  the  blast.  On    the 
the  other  hand  the  reaction 


takes  place  rapidly  even  at  relatively  low  temperatures.  More- 
over, metallic  Ni  alloys  very  readily  with  its  sulphides  and  with 
NiO. 

Reaction  smelting,  therefore,  yields  under  the  most  favorable 
conditions  only  an  alloy  consisting  of  much  NiO  with  little 
Ni;  it  can  never  lead  to  metallic  nickel  under  any  of  the 
methods  of  working  which  have  been  tried  up  to  the  present 
time. 

Desulphurizing  Fusion  of  Concentrated  Matte.  Borchers 
and  Lehmer's  process  consists  of  smelting  the  concentrated 
matte  with  lime  and  carbon,  preferably  in  an  electric  furnace. 
If  the  matte  was  entirely  free  from  other  metals,  then  this 
smelting  can  be  carried  out  so  that  a  fused,  pure  nickel  will 
be  obtained  directly.  Since,  however,  the  sulphide  nickel  ores 
usually  contain  precious  metals  (Au  and  members  of  the  Pt 
group),  the  product  is  an  alloy  of  Ni  with  these  metals, 
from  which  pure  nickel  is  obtained  by  electrolysis  (see  Nickel 
Refining).  The  slag  of  CaS  can  be  used  either  as  a  sulphide 
flux  in  the  smelting  of  matte,  or,  as  in  the  above-described 
process  of  Borchers  and  Warlimont,  as  a  precipitant  for  Co 
and  Ni  solutions. 


(C)    Nickel  Refining 

As  a  raw  material  for  the  preparation  of  pure  nickel,  the  crude 
Ni  prepared  as  described  under  B  is  usually  taken.  It  is  pos- 
sible, however,  to  make  use  of  concentrated  matte,  or  even 
certain  ores,  as  starting  material. 


172  METALLURGY 

Pure  Nickel  from  Crude  Nickel.  The  crude  nickel  obtained 
by  a  reducing  roast,  or  the  nickel  obtained  by  a  reducing  smelt 
of  nickelous  oxide,  contains,  as  has  already  been  mentioned, 
C  and  NiO  so  that  it  is  not  suitable  to  be  used  in  the  manu- 
facture of  nickel  wares  (foil,  wire,  etc.). 

Refining  Fusion.  The  removal  of  the  C  offers  no  difficulties, 
as  it  takes  place  when  the  crude  metal  is  melted ;  but  the  last 
traces  of  NiO  cannot  be  removed  by  means  of  carbon.  This 
reduction  of  the  NiO  was  first  accomplished  by  the  method 
of  Fleitmann,  who  fused  with  Mg,  which  is  an  extremely 
energetic  reducing  agent;  then  Basse  and  Selve  accom- 
plished the  same  end  by  mixing  in  as  much  as  3%  of 
MnO2  before  carrying  out  the  reducing  roast.  The 
MnO2  is  reduced,  together  with  the  NiO,  and  serves 
to  remove  the  NiO  when  the  cube  nickel  is  heated  in 
carbon-free  crucibles;  naturally  it  prevents  the  forma- 
tion of  NiO  by  atmospheric  oxygen  when  the  metal  is 
melted. 

Electrolysis  is  employed  if  the  nickel  contains  precious  metals. 
Since  the  solution  tension  of  Ni  is  on  the  other  side  of  H, 
the  concentration  of  H  ions  must  be  kept  as  low  as  possible, 
contrary  to  the  practice  in  the  electrolytic  refining  of  Cu. 
This  is  accomplished  by  limiting  the  amount  of  free 
acids  having  high  dissociation  constants  (HC1,  H2SO4)  or 
by  using  acids  with  low  dissociation  constants  (H3BO3, 
HsPO4,  and  organic  acids).  Unfavorable  changes  in  con- 
centration are  avoided  by  moving  the  electrodes  and  heating 
the  electrolyte.  When  these  requirements  are  fulfilled,  the 
electrolytic  deposition  of  Ni  offers  no  difficulties  worth 
mentioning.  The  conditions  for  carrying  out  the  elec- 
trolysis are  as  follows: 

ELECTROLYZING  TANKS:    made  of  wood  with   lead   lining. 
ANODES:  Cast  plates  of  impure  nickel,  containing  precious 

metals  with  S  up  to  3%. 
CATHODES  :  Pure  Ni. 
ELECTROLYTE:  NiSO.j  or  NiCU  solution,  the  former  being 


NICKEL  173 

preferred.     Concentration  4  to   14  oz.  Ni  to  the  gallon 
not  under  0.01%  or  over  0.25%  of  free  acid. 
TEMPERATURE:  50°  to  90°  C.  (122°  to  194°?.) 
CURRENT  DENSITY:  5  to  28  amperes  per  sq.ft. 
POTENTIAL:    i  to  1.3  volts  with  15  or  20  amperes  per  sq.ft. 
REACTIONS  DURING  THE  ELECTROLYSIS:    under  the  above 
conditions  it  is  possible,  even  when  the  crude  metal  con- 
tains up  to  0.5%  Cu,  to  deposit  a  sufficiently  pure  Ni 
(0.1%  to  0.2%  Cu).     It  is  important  to  have   enough   S 
present  not  only  to  precipitate  the  Cu  as  Cu2S,  but  also 
to  make  the  last  traces  of  Fe  form  compounds  of  Cu2S 
with   FeS.     Such  compounds  are  attacked  by  the   elec- 
trolysis less  in  proportion  as  the  current  density  at  the 
anode  is  kept  low,  i.e.,  the  more   the  anode    surface   is 
increased  in  proportion  to  the  size  of  cathode  surface.    In 
designing  an  electrolyzing  plant  for  Ni,  therefore,  the  anodes 
should  be  made  relatively  large  and  the  cathodes  small. 
Pure  Nickel  from  Concentrated    Matte.     The    matte  which 
is  obtained  as  end  product  in  the  concentration  of  sulphide 
ores  contains  up  to  76%  Ni.    It  then  has  23%  to  24%  S,  up  to 
0.4%  Fe,  0.1%  to  0.2%  Cu,  and  small  amounts  of  SiO2  due 
to  the  presence  of  a  small  amount  of  enclosed  slag.     Accord- 
ing, to  the  above-cited  researches  of  Bornemann,   unlike  the 
concentrated  matte  of  copper    smelting,  it  does  not  contain 
a  pure  sulphide  of  nickel,  but  rather  a  solution  of  Ni  in  NisS2. 
This  matte  already  has  many  metallic  properties  and  is  suited, 
as  has  been  demonstrated  by  Borchers  and   Giinther,  for  a 
Direct  Electrolytic  Treatment.     The  conditions,  as  established 
by  Giinther  are  partly  the  same  as  in  the   electrolysis  of 
crude  Ni.    The  points  of  difference  may  be  summarized  as 
follows : 

ANODES:  concentrated  matte,  cast  into  plates. 
CURRENT  DENSITY:  23  to  26  amperes  per  sq.ft. 
POTENTIAL:  3  volts. 

REACTIONS  DURING  THE  ELECTROLYSIS:    Of  the  constituents 
of  the  matte,  it  is  chiefly  Ni  that  passes  into  solution  at 


174  METALLURGY 

the  anode,  leaving  behind  free  S,  which  remains  adhering 
to  the  electrode  as  a  coherent  but  porous  crust  until  the 
matte  has  nearly  disappeared,  and  thus  influences  the 
progress  of  the  electrolysis  scarcely  at  all.  In  this  crust 
there  are  present  about  80%  free  S  and  20%  of  a  mixture 
of  Cu,  Fe  and  Ni  embedded  in  it.  Whereas  the  original 
.  matte  contained  less  than  0.2%  Cu,  the  mixture  of  embedded 
sulphides  will  contain  over  12%  Cu  with  about  51%  Ni 
and  3.5%  Si  +  C.  From  this  composition,  and  from  the 
low  Cu  content  of  the  Ni  that  is  deposited  on  the  cathode, 
it  follows  that  Cu  is  combined  with  S  in  the  crust  and 
that  the  latter  unites  with  the  sulphides  of  Fe  and  Ni  to 
remain,  for  the  most  part,  adhering  at  the  anode;  this  is 
confirmed  by  the  results  obtained  in  the  electrolysis  of 
crude  Ni  that  contains  S. 

Electric   Smelting   of   Concentrated   Nickel   Matte   with  lime 
and  carbon  directly  yields,  as  already  mentioned  under  B, 
pure  Ni  provided  the  matte  contains  no  foreign  metal  other 
than  S  and  Ni. 
Pure  Nickel  Directly  from  Ores. 

Extraction  of  Ni  by  Means  of  CO,  Mond's  Process.  This  is  based 
upon  the  property  that  Ni  possesses  of  combining  with  CO 
to  form  a  readily  volatile  compound  called  nickel  carbonyl, 
Ni(CO)4.  The  process  is  applicable  only  to  ores  and  roasted 
products  that  contain  the  Ni  in  the  form  of  free  oxides.  The 
operations  are  as  follows: 

1.  REDUCING  ROAST  in  retorts  heated  to  300°  C.  with  gases 
containing  H.     The  NiO  is  changed  by  this  treatment 
into  porous  Ni.     It  is  important  to  maintain  a  reducing 
temperature  as  low  as  possible  so  that  a  large  surface  of 
the  reduced  Ni  will  be  obtained. 

2.  PASSING  OF  GASES  CONTAINING  CO  at  100°  C.  and  15 
atmospheres  pressure  over  the  charge.     The  temperature 
of  100°  C.  is  chosen  because  it  can  be  kept  constant  easily 
by  means  of  steam.     As  regards  the  pressure,  Dewar  found 
that  Ni(CO)4  is  stable  at 


NICKEL  175 

50°  under  a  pressure  of     2  atmospheres. 
100°        "  "          15 

180°        "  "          30 

250°        "  "        100 

3.  Deposition  of  Ni  from  the  Ni(CO)4,  which  depends  upon 
the  ready  reversibility  of  the  reaction 


The  vapors  of  Ni(CO)4  escaping  from  the  vessels  under 
pressure  can  be  decomposed  into  Ni  and  CO  either 
directly,  or,  after  a  preliminary  purification  which  con- 
sists in  cooling  and  freeing  from  mechanical  admix- 
tures, by  simply  heating  at  200°  C.  under  ordinary 
atmospheric  pressure. 

The  CO  is  collected  and  used  again  in  the  process. 
Properties  of  Nickel  : 

SPECIFIC  GRAVITY:  9. 

COLOR:  light  gray. 

MECHANICAL  PROPERTIES  :  possesses  a  high  degree  of  ductility 
and  tenacity  so  that  it  can  be  rolled  into  foil  and  drawn 
into  wire. 

STRUCTURE  of  Cast  Ni:  similar  to  that  of  ferrite  (see  iron); 
rolled  Ni  is  very  fine  grained. 

MELTING  POINT:   1451°  C.  (2644°  F.). 

THERMAL  AND  ELECTRICAL  CONDUCTIVITY:  about  0.2  that 
of  Ag. 

ALLOYS:  with  most  metals.  Important  alloys  are  coin 
metal,  Niand  Cu;  German  silver,  Ni,  Cu  and  Zn;  nickel 
steel,  Fe  and  Ni.  In  the  extraction  of  Ni  the  tendency 
to  alloy  with  its  own  sulphides  and  arsenides,  as  well  as 
with  NiO,  comes  into  consideration. 

CHEMICAL  BEHAVIOR:  The  metal  resists  oxidation  fairly  well 
at  moderate  temperatures.  Waste  in  forging  and  rolling 
is  slight.  Solution  tension  in  acids  and  salts  is  greater 


176  METALLURGY 

than  H( +0.223  vo^ts  toward  H).  It  dissolves,  therefore, 
quite  slowly  in  HC1  and  H^SC^,  more  rapidly  in  HNOa, 
forming  under  normal  conditions  a  bivalent  Ni  cation. 
In  the  molten  condition  the  solution  tension  is  greatest 
in  arsenides,  next  in  sulphides;  in  arsenides  the  solution 
tension  is  greater  and  in  sulphides  less  than  that  of  Cu. 


IRON 

Sources 

Natural  Sources  : 

NATIVE,  as  a  constituent  of  the  earth's  crust  it  is  of  rare 

occurrence;    large  lumps  of  meteoric  iron  are,  however, 

sometimes  found  (up  to  55,000  Ibs.  in  weight). 
As   OXIDE  AND  HYDRATED  OXIDE  in  the  following  ores: 

Specular  Iron  Ore  l 

Hematite  }•  Fe2O3  with  70%  Fe. 

Red  Iron  Ore 

Magnetite,    or    Magnetic    Iron    Ore:      FeO.Fe2O3    with 
72.4%  Fe. 

Brown  Iron  Ore      j 

Limonite,  Bog  Ore  !>  Fe2O2(HO)2  to  Fe2O(OH)4. 

Brown  Hematite      J 
As  SULPHIDE  in 


M  " 

Marcasite 

Magnetic  Pyrites,  or  pyrrhotite,  FeS  with  some  FeS2, 
usually  nickeliferous.  Combined  and  mixed,  as  FeS 
with  FeS2,  in  numerous  other  sulphides. 

The  sulphides  themselves  do  not  come  into  considera- 
tion as  ores  of  iron  until  after  they  have  been  utilized 
for  the  manufacture  of  H2SO4j  they  are  then  roasted 
to  oxides  so  that  they  retain  but  very  small  amounts  of 
sulphide.  (Roasted  or  "  burnt  "  pyrite,  see  below). 
SALTS. 

Spathic  Iron  Ore,  or  Siderite,  FeCO3  with  48.27%  Fe, 
usually  carrying  some  MnCOs,  CaCOy  and  clay. 

Blue    Iron  Ore,    or    Vivianite,  Fe3(PO4)2,  which  seldom 

177 


178  METALLURGY 

occurs    by   itself,    but   accompanies   other   ores     (e.g., 
limonite). 
Other  Sources : 

METALS:    Scrap  from  the  mechanical  working  of   iron  and 

old  metal. 

OXIDES:   Roasted  pyrite,  the  roasted  residue  from   sulphide 

ores.     Hammer   scale,    the   oxide   crust   formed    on    iron 

heated  to  redness  and  detached  in  the  mechanical  working. 

SALTS:  Basic  silicates  and  phosphates,  slags  from  the  refining 

of  iron. 

(A)   Concentrating  and  other  Preliminary 
Operations. 

For  Concentrating  Iron  Ores  it  is  often  desirable  to  employ: 
Magnetic  Concentration  which  is  carried  out  largely  when  the 
ore  is  magnetite.     Ores  carrying  hematite  are   also  concen- 
trated by  the  magnet,  after    they  have  been  converted   into 
magnetite  by  roasting. 

The  Briquetting  of  pulverulent  ores  and  other  raw  material  has 
been  carried  out  to  a  considerable  extent  in  recent  years.     The 
requirements  of  a  good   ore  briquette  for  blast  furnace  work, 
have  never  been  satisfied. 
Chemical  Preparation. 

Dissociating  Roast,  usually  combined  with  an  oxidizing  roast, 
is  carried  out  to  effect  the 
DEHYDRATION  of  hydrated  ores  (limonite). 
OXIDATION  of  magnetite. 
DECOMPOSITION  AND  OXIDATION  of  carbonates  (spathic  iron 

ore). 

WORKING  up  of  pyrite  for  sulphuric  acid. 
The  Apparatus  used  for  this  work  consists  of  simple  shaft 
furnaces,  the  so-called  calcining  furnaces  which  usually  work 
with  natural  draft.  They  either  take  both  ore  and  fuel,  or 
the  ore  alone,  when  they  are  fired  externally  with  solid  or 
gaseous  fuel.  The  discharge  is  usually  through  a  saddle- 


IRON 


179 


180 


METALLURGY 


shaped  bottom  to  facilitate  the  taking  out  of  the  solid 
charge  through  side  openings.  The  apparatus  that  serves 
specially  for  the  roasting  of  pyrite  has  been  described 
under  Copper. 


FIG.  166.— Old  Siegen  Kiln 

(B)   Cast  Iron 

Reducing  Roast.  Various  attempts  have  been  made  to  make 
use  of  this  operation  and  produce  the  so-called  "  iron  sponge," 
a  product  which  is  metallic,  but  is  not  fused  until  it  comes  to 
the  refining  operation.  Such  processes  have  never  been  adopted 
permanently,  even  for  the  preparation  of  iron  for  special  pur- 
poses in  which  it  was  thought  desirable  to  have  the  iron  in 
this  form. 

The  Reducing  Smelt  is  to-day  the  only  process    used  in    the 
iron  industry  for  the  production  of  cast  iron. 
As  Reducing  Agent,  carbon  in  the  form  of  charcoal  is  seldom 

used,  but  rather  coke  and  the  CO  formed  in  the  furnace. 
The  Other  Fluxes  are  determined  not  alone  by  the  nature  of 
the  gangue  to  be  removed,  but  by  the  purpose  which  it  is 
desired  to  accomplish,  and  by  the  demands  that  are  to  be 


IRON  181 

placed  upon  the  iron.  Thus  for  slagging  off  the  SiC>2  and 
clay  which  is  frequently  present  in  the  ore,  CaO  in  the  form 
of  CaCC>3  is  added  to  the  charge,  and  the  amount  added 
depends  upon  whether  it  is  desired  that  the  cast  iron  pro- 
duced shall  have  little  or  much  Si.  To  produce  a  cast  iron 
rich  in  Si,  it  is  necessary  to  form  a  more  basic  silicate  and  a 
slag  richer  in  AloOs  than  when  an  iron  low  in  Si  is  desired. 
For  slagging  off  the  S  from  ores  containing  some  sulphide, 
the  slag  should  be  kept  as  basic  as  possible  with  CaO.  The 
addition  of  an  ore  rich  in  Mn  can  also  serve  to  advantage  in 
this  case. 

For  Fuel,  coke  is  generally  used  and  charcoal  but  seldom. 
In  some  few  cases  where  coke  is  expensive  and  the  ore  deposits 
are  in  the  vicinity  of  cheap  water  power,  electricity  is  now 
to  be  considered  as  a  possible  source  of  heat.  The  blast 
requisite  for  the  production  of  the  desired  amount  of  heat 
by  the  combustion  of  carbon  is  given  a  preliminary  heating 
of  750°  to  900°  C.  and  compressed  up  to  one  atmosphere. 
Apparatus : 

BLAST  FURNACES.  These  shaft  furnaces  are  sometimes  as 
much  as  TOO  feet  in  height.  The  shaft  rests  upon  pillars 
and  is  built  of  masonry  with  iron  bands,  or  of  an  iron 
casing,  and  has  an  inside  diameter  of  26  ft.  at  the 
widest  part,  narrowing  toward  the  throat,  which  has  about 
two-thirds  this  width.  The  iron  blast  furnaces  usually 
have  internal  crucibles.  The  hearth  has  an  inside  diam- 
eter of  13  ft.  and  a  maximum  height  of  nj  ft.  It  contains 
the  tuyeres  for  the  blast,  the  cinder  notch,  and  the  tap  hole. 
One  tuyere  is  reckoned  for  each  sq.  yd.  of  hearth  surface 
so  that  with  a  diameter  of  12  ft.  there  are  n  or  12  tuyeres. 
These  tuyeres  are  hollow  bronze  or  copper  castings  which 
can  be  cooled.  The  boshes  rest  upon  the  hearth  and 
reach  into  the  shaft  to  about  two-fifths  of  the  whole  height 
of  the  furnace,  reckoning  from  the  bottom  of  the  hearth, 
and  they  widen  as  they  go  upward  until  they  reach  the 
widest  part  of  the  furnace.  The  boshes  and  hearth  are 


182 


METALLURGY 


both  cooled  by  spraying  with  water.  Large  furnaces 
use  about  525  gallons  of  water  per  minute.  Iron  blast 
furnaces  work  with  a  closed  throat,  for  the  hot  escaping 
gases  are  rich  in  CO  and  can  be  utilized  as  a  source  of 
heat  and  power.  The  "down-comer"  is  usually  a  wide 


•irtnnnrtntlTn 

~U     LJ    L.J    LJ    L      LJ     LJ     [J  LJ 


FIG.  167. — Blast  Furnace  with  Stoves  (Friedrich-Alfred  Smelter)  Rheinhausen. 

central  sheet-iron  tube  passing  through  the  bell  and  hopper- 
feed  in  the  throat.     In  some  constructions,  the  gas  is  taken 
off  laterally  from  the  closed  throat  of  the  furnace. 
HOT    BLAST.     To-day    regenerator    stoves    of    about    the 
same  dimensions  as  the  blast  furnaces  usually  form  an 


IRON 


183 


inseparable  whole  with  the  latter  throughout  all  of 
their  operations.  The  regenerative  system  requires  the 
setting  up  of  at  least  three  stoves;  to  be  on  the  safe 
side  it  is  customary  to  allow  four  stoves  for  each  blast  fur- 
nace. In  the  cylindrical  stoves  of  Cowper's  design,  there 


FIG.  168. — Blast  Furnace  with  Dust  Catcher  and  Blast  Stove. 

is  found  a  shaft  into  which  the  hot  gases  from  the  throat  of 
the  blast  furnace  enter  at  the  bottom  for  the  CO  to  be 
burned  by  air  which  has  been  heated  by  passing  through  flues 
in  the  walls  of  the  shaft.  The  remaining  space  of  the  regen- 
erator is  loosely  filled  with  special  brick,  the  spaces  between 
which  form  numerous  little  channels.  It  forms  the  real 


181 


METALLURGY 


FIG.  170 


F.G.169 


FIG.  172. 


FIG.  173. 


IRON  185 

heat  reservoir.  The  products  of  combustion  are  divided 
into  innumerable  gas  currents  by  these  brick,  and  in  this 
part  of  the  shaft  they  are  carried  downward  and  then  led 
away  from  the  open  space  at  the  bottom.  After  the  stove 
has  become  hot  enough,  the  gas  and  combustion  air  are 
led  into  a  neighboring  regenerator,  and  now  the  air  intended 
for  the  blast  furnace  is  passed  through  the  hot  chamber  in 
the  opposite  direction  to  that  previously  taken  by  the 
gases  from  the  blast  furnace.  This  heated  air  collects 
above  the  part  of  the  regenerator  that  is  filled  with  brick 
and  passes  downward  through  the  combustion  chamber 
and  thence  into  the  main  pipe  for  the  blast.  Since  this 
air  must  be  heated  to  between  750°  and  900°  C.,  the  iron 
pipes  through  which  it  passes  are  provided  with  a  fire- 
brick lining.  Even  the  bustle  pipe  leading  to  the  sep- 
arate tuyeres  of  the  blast  furnace  is  lined  or  else  placed 
in  a  hollow  tube  with  non-conducting  layer  of  air  inter- 
vening. 

REACTIONS  IN  THE  BLAST  FURNACE.  The  changes  within 
the  more  important  temperature  boundaries,  which  the 
solid  charge  experiences  as  it  passes  from  the  throat  of 
the  furnace  downward,  are  the  following:  With  furnaces 
that  are  75  to  100  ft.  high  between  the  throat  and  the 
bottom  of  the  hearth,  the  temperature  at  the  top  ranges 
between  150°  and  300°  C.,  whereas  in  the  smelting  zone 
the  temperature  may  reach  1500°. 

150°  to  400°  C. 

Drying  of  the  charge.  Breaking  down  of  hydrates.  Begin- 
ning of  reducing  reactions. 

400°  to  1000°  C. 

According  to  the  studies  of  Boudouard  on  the  equilibrium 
of  the  reaction 


186 


METALLURGY 


the  speed  of  the  reaction  in  the  direction  of  the  upper 
arrow  increases  from  about  400°  to  700°,  but  from  there 
on  the  tendency  for  the  reaction  to  take  place  in  the  opposite 
direction  becomes  more  and  more  pronounced.  Here 
the  efficiency  of  CO  as  a  deducing  agent  rapidly  diminishes. 
As  regards  the  concentration  of  the  reaction  gases,  it  has 
been  found  that  the  ratio  of  CO:CO2  must  not  be  less 


25m. 


As  reducing  agent 


FIG.  174. 


FIG.  175. 


than  i :  i  at  the  most  favorable  reaction  temperature  (700°). 
At  such  a  concentration  of  CC>2,  the  reducing  action  of 
the  CO  stops  almost  entirely  even  at  700°  C. 

In  this  zone,  therefore,  the  greater  part  of  the  oxides 
of  iron,  namely  Fe2O3,  FesO4,  and  FeO,  are  reduced  by 
CO,  but  at  about  800°  there  is  a  dissociation  of  the  car- 
bonates, CaCO3,  FeCOs  and  MnCO3,  into  CaO,  FeO, 


IRON  187 

MnO,  and  CC>2,  and  the  effect  of  this  is  to  greatly  increase 
the  concentration  of  the  CO2- 

From  1000°  upward 

the  C  begins  to  dissolve  in  solid  Fe.  Disregarding  other 
influences  (Si,  Mn,  etc.)  it  has  been  established  that  the 
solubility  of  C  in  Fe  increases  as  the  temperature  rises 
(cf.  also  p.  190). 

From  1300°  upward 

solid  C  is  practically  the  only  reducing  agent,  particularly 
for  FeO,  MnO,  SiC>2,  phosphates,  and  silicates;  FeS  also 
reacts  with  CaO  and  C  to  form  Fe  and  CaS. 

The  blast,  as  it  passes  through  the, furnace  in  the  opposite 
direction,  of  course  loses  its  O  rapidly  by  reason  of  the 
combustion  of  the  C.  The  more  the  blast  has  been  heated 
before  it  comes  in  contact  with  the  coke,  the  greater  will 
be  the  formation  of  CO(C+-€^  =  CO2  and  CO2  +  C  =  2CO). 
The  CO,  after  it  is  formed,  as  the  above  equations  show, 
will  be  changed  repeatedly  into  CO2  and  back  again  to 
CO.  The  gas  finally  leaving  the  throat  of  the  furnace 
has  a  composition  of  22  to  28%  CO,  16  to  8%  of  CO2, 
63  to  57%  N,  and  a  low  per  cent  of  hydrocarbons  and  free 
hydrogen. 

With  the  present-day  dimensions  of  blast  furnaces, 
about  500  tons  of  pig  iron  are  produced  in  a  furnace  during 
a  day  of  24  hours.  The  weight  of  a  single  charge  of 
ore  (ore  and  slagging  fluxes)  reaches  15  tons/ for  which 
3  to  7  tons  of  fuel  are  required.  The  charge  remains 
in  the  furnace  for  from  18  to  26  hours.  Aside  from  the 
content  of  the  ore  and  the  quality  of  the  other  constituents 
of  the  charge,  the  capacity  of  the  furnace  is  dependent 
particularly  upon  the  nature  of  the  pig  iron  that  it  is 
desired  to  produce.  A  furnace  capable  of  producing 
100  tons  of  white  cast-iron,  can  yield  at  the  most  but  80 
tons  of  gray  iron,  and  still  less  of  other  kinds,  e.g.,  metal  run- 


188  METALLURGY 

ning  high  in  Mn  or  Si.  Furthermore,  in  respect  to  the  con- 
sumption of  fuel,  the  production  of  white  iron  is  the  most 
favorable.  For  such  iron,  100  to  no  tons  of  coke  are 
required  to  produce  100  tons  of  the  metal.  The  coke 
consumption  rises  with  other  qualities  of  iron.  Concern- 
ing the  nature  and  value  of  the 
Products  of  the  Blast  Furnace,  we  may  summarize  them  as 

follows : 

i.  PIG  IRON,  an  alloy  of  Fe  and  compounds  of  Fe  and  Mn 
with  C,  Si,  P,  S,  and  O,  from  which  on  solidifying  individual 
constituents,  such  as  C  in  the  form  of  graphite  or  amorphous 
carbon,  carbides,  silicides,  etc.,  may  separate  out.  The 
total  C  content  is  over  2%  and  usually  under  5%. 

The  nature  of  the  ore,  the  conditions  of  working,  and 
the  qualities  required  by  the  purchaser,  have  led  to  the 
production  of  quite  a  number  of  different  commercial 
grades  of  pig  iron,  but  the  qualities,  as  in  the  case  of  most 
pure  irons,  arejdependent  largely  upon  the  amount  and 
jnanner  of^com^ina^iQiL-QLthejC  present.  The  latter^  on" 
its  part,  is  influenced  by  other  constituents  found  in  the 
iron,  particularly  Si  and  Mn;  a  short  discussion  of  the 
relations  between  Fe  and  C  will  explain  best  the  nature 
of  the  different  products  obtained  in  the  iron  industry. 
If  we  start  with  a  perfectly  liquid  melt,  it  is  known  from 
the  interesting  researches  of  Roberts- Austen,  Wiist  and 
Goerens,  and  others,  that  all  the  C  can  exist  combined 
with  Fe  as  Fe3C  (which  is  called  cementite)  until  the 
amount  of  C  present  reaches  6.66%.  A  higher  C  con- 
tent scarcely  comes  into  consideration  in  practical  work; 
with  a  lower  C  content  there  may  exist  in  the  fusion  a 
solution  of  this  carbide  in  Fe.  If  the  temperature  falls, 
it  may  happen  when  the  amount  of  Fe3C  present  is  large 
that,  just  before  the  whole  melt  solidifies,  a  decomposition 
of  the  Fe3C  into  3Fe  +  C  takes  place. 

Directly  after  such    decomposition  free  C  is  therefore 
embedded  as  pure  graphite  in  pure  Fe,  and  as  the  melting- 


IRON  189 

point  of  the  latter  is  higher  than  that  of  the  solution  (about 
1500°  C-)  it  begins  to  solidify  within  the  liquid  melt  and 
thereby  envelopes  the  solid  graphite.  As  long  as  the  lat- 
ter possesses  this  specifically  heavy  shell  it  remains  uni- 
formly distributed  in  the  melt.  When,  therefore,  the  cool- 
ing does  not  takeplace  toojslowly,  a  pig  iron  is  obtained 
^hich  containsTarge  crystals  of  graphite  uniformly  dis- 
tributed through  it.  When,  however,  the  temperature 
is  kept  high  for  a  long  time,  the  crust  of  Fe  gradually 
becomes  saturated  again  with  C,  becomes  liquid  and  the 
graphite  now  rises  to  the  surface  and  forms  a  scum  called 
"  kish."  The  melt,  in  which  all  this  has  taken  place, 
finally  solidifies  on  further  cooling  at  1130°  C.,  with  about 
4.2%  C  chemically  combined  with  Fe. 

From  molten  iron  with  less  than  4.2%  C  and  more  than 
2%  C,  crystals  first  separate  from  the  melt  which  are  to 
be  regarded  as  solid  solutions  of  Fe3C  in  Fe  (mixed  crys- 
tals). The  saturated  solution,  or  in  other  words  the  mixed 
crystals  richest  in  C,  contain  2%  of  C,  they  correspond, 
therefore,  to  a  solid  solution  of  about  2  Fe3C  +  i5Fe. 

The  mother  metal  from  which  these  crystals  are  deposited 
thus  becomes  enriched  in  Fe3C  until  the  above-mentioned 
4.2%  C  content  is  reached.  Then  the  remaining  metal 
solidifies  at  1130°  C. 

At  the  concentration  4.2%  C,  lies  a  eutectic  point  between 
the  melting  points  of  the  saturated  mixed  crystals,  2Fe3C  -f- 
i5Fe  (2%  Fe)  and  cementite,  Fe3C  (6.66%  Fe),  and  this 
eutectic  corresponds  closely  to  the  composition  Fe3C  +  Fe2. 
As  Wtist  and  Charpy  both  showed  independently  in  1905, 
this  eutectic  solidifies  under  all  conditions,  no  matter 
what  other  substances  may  be  present,  at  1130°  C. 

Under  the  conditions  corresponding  to  the  line  aB  (Fig. 
176),  there  exists  a  solid  mass  of  free  mixed  crystals  em- 
bedded in  eutectic;  at  the  point  B  there  is  only  eutectic 
present;  and  at  the  line  BC  graphite  and  cementite  in 
eutectic. 


190 


METALLURGY 


In  every  cast  iron  with  more  than  2%  C,  there  tends  to 
take  place  after  solidification  a  decomposition  of  Fe3C  into 
3Fe  +  C  (graphite  or  amorphous  carbon,  temper  carbon) 
provided  the  Fe3C  has  already  crystallized  from  the  solution 
or  has  had  an  opportunity  to  crystallize.  Substances  which 
favor  the  crystallization  of  this  compound  from  the  liquid 
or  solid  solution  (e.g.,  Si,  Al?)  naturally  favor  as  well  the 
progress  of  the  reaction  Fe3C  =  3Fe  +  C;  substances 


1500 
1400 
1300 
1200 
1100 
1000 
900 

800 
M 

A 

\ 

~^ 

^ 

^ 

] 

jiqui< 

I 

\ 

Liq 

lid 

^> 

\ 

•S 

/ 

N5 

ixedC 

I 
V 

£S 

:3C 

\ 

\ 

/ 

/ 

Graj 

hite 

Mix 
I 

edCrj 

en+F< 

stals 
)3C 

5 

§/ 

f 

+  ll 

quid 

^ 

Mixlci.  ^rsst. 

H-Knt. 

Kl 

Hectic 

.  Re, 

O-j-T? 

lit. 

/ 

/r 

G 

/ 

Mixe 

ICrys 

tals   - 

1-  Gra 

phite 

4-Te 

mper 

^arbo 

i 

\ 

/ 

O^ 

x/ 

VOO 
600 

500 

1 

Ft 

rrite 

\-  Gri 

phite 

-fTe 

mper 

Darbo 

v 

i 

0.5      1,0       1,5       2,0       2,5       3,0       3,5       4,0       4,5       5,0       5,5  #  C. 

FIG  176. 

which  increase  the  solvent  power  of  Fe  for  Fe3C  (e.g.,  Mn) 
increase  the  stability  of  the  FeaC  and  hence  tend  to 'prevent 
the  formation  of  free  C.  From  this  point  of  view  the  nature 
of  the  two  principal  kinds  of  cast  iron  appears  in  the  follow- 
ing light: 

WHITE  IRON  is  a  supercooled  solution  and  is,  therefore, 
to  be  regarded  as  representing  a  meta-stable  system 
between  Fe3C  and  Fe,  in  which  the  reaction  Fe3C  = 


IR©N  191 

3Fe  +  C  has  not  been  allowed  to  take  place.  The  chilled 
white  iron  consists  of  cementite  and  pearlite  (see  Fig. 
178),  also  containing,  when  the  cooling  was  very  rapid, 
some  mixed  crystals  of  the  concentration  FeaC :  Fen. 

If  such  an  iron  is  kept  for  a  long  time  at  about  1000°  C. 
the  compound  FesC  begins  to  crystallize  from  the  solid 
solution  and  then  the  Fe3C  decomposes  rapidly,  although 
the  free  C  does  not  appear  as  graphite,  but  as  the  amor- 
phous, readily-oxidizable,  temper  carbon  (see  Fig.  180, 

P-  195)- 
GRAY  IRON  represents  the  more  stable  system  Fe — Fe3C — C. 

It  has  had  time,  at  the  different  temperatures  and  con- 
centrations, to  reach  more  or  less  completely  a  state 
of  equilibrium,  i.e.,  during  the  cooling  FeaC  has  had 
opportunity  to  crystallize  out  and  dissociate  into  3Fe 
and  C  and  the  latter  is  now  to  be  found  in  the  form  of 
graphite,  although  it  is  not  impossible  that  some  temper 
carbon  may  have  been  formed  in  the  vicinity  of  1000°. 
(Fig.  179.)  As  has  been  pointed  out,  the  separation 
of  C  is  favored  by  the  presence  of  Si,  even  when  the 
C  content  is  low,  and  to  promote  this  separation  from 
i  to  2.5%  Si,  or  even  5%  Si  in  special  cases,  is  often 
added  to  the  iron. 

*  Whereas  gray  iron  is  used  for  foundry  purposes  as  well  as 
for  the  production  of  wrought  iron,  white  iron  is  used  almost 
exclusively  for  the  latter  purpose.  A  distinction  is  further 
made  between  foundry  iron  and  Bessemer  iron  (Si  high, 
P  low),  open  hearth  iron,  basic  Bessemer  iron  (high  in  Mn 
and  P),  spiegeleisen  or  ferromanganese  (rich  in  Mn,  low 
in  P),  and  charcoal  iron  (rich  in  C). 

In  the  many  constituents  of  the  different  kinds  of  pig 
iron,  only  a  few  of  the  particularly  important  solidification 
phenomena  have  been  explained.  The  first  accurate 
equilibrium  diagrams  between  Fe  and  C  were  those  of 
Roberts-Austen  and  of  Roozeboom.  Subsequently  Heyn 
gave  different  diagrams  for  the  meta-stable  and  stable 


192  METALLURGY 

systems,  while  Benedicks  improved  the  original  diagram 
of  Roozeboom.  Since  Wiist  and  Charpy  have  determined 
the  freezing-point  of  the  system  Fe— Fe3C  to  be  1130° 
for  all  cases,  and  Wiist  and  Goerens  have  proved  that 
graphite  is  always  produced  by  the  decomposition  of 
Fe3C,  Goerens'  diagram  (Fig.  176)  may  be  taken  as 
correct  for  the  stable  and  for  the  super-cooled  system 
Fe  -  Fe3C.  To  understand  the  terms  used  in  metallography, 
the  following  definitions  will  serve: 


FIG.  177. — Ferrite  (X5oo). 

FERRITE,  chemically  pure  iron:  a-iron,  which  is  magnetic 
and  free  from  C,  passes  at  780°  into  /?-iron,  which  is 
non-magnetic  and  also  practically  incapable  of  dissolving 
C.  This  second  form  of  iron  exists  only  between  780 
and  800°  C.  Above  880°  it  changes  to  ^-iron,  which 
is  likewise  non-magnetic,  but  is  capable  of  dissolving 
C  or  Fe,C  (Fig.  177). 

CEMENTITE  is  iron  carbide,  Fe3C  (Fig.  178). 

AUSTENITE  and  MARTENSITE  are  solid  solutions  of  Fe3C 
in  f-iron.  They  exist  as  mixed  crystals  of  varying  con- 
centrations. The  saturation  point  of  these  solutions 
lies  at  2%  C  corresponding  to  the  composition  2Fe3C  -f 


IRON  193 

i5Fe.     A  eu  Lectio  between  these  saturated  mixed  crystals 
and  Fe3  C  lies  at  4.1%  C,  corresponding  to  a  composition 


TROOSTITE:  a  colloidal  solution  of  Fe3C  in  Fe. 

SORBITE:    mixtures  of  Fe,  Fe3C,  and  solid  solutions  of 

Fe3C  in  Fe. 
PEARLITE:  the  eutectic  between  ferrite  (Fe)  and  cementite 

(Fe3C).    Its  C  content  is  0.9%,  corresponding  to  Fe3C  + 

2oFe  (Figs.  178  to  180). 


FiG.  178. — White  Cast  Iron.     Cementite  (White)  with  Lamellar  Pcarlite  (X5<x>). 

GRAPHITE  (Fig.  179):  concerning  its  formation  see  p.  188. 
TEMPER  CARBON  is  non- graphitic  C  which  separates  out 
from  white  iron  by  keeping  it  for  a  long  time  at  a  tem- 
perature near  1000°,  during  which  time  the  finely-divided 
cementite  changes  into  a  mixture  of  ferrite,  pearlite  and 
temper  carbon  (Fig.  180). 

This  form  of  carbon  is  more  readily  oxidizable  than 

graphite  or  carbide  carbon  (cf.  Malleableizing,  p.  198). 

2.  SLAG.     SiO2  and  A12O3  are  almost  invariably  present  in  the 


194  METALLURGY 

gangue  of  iron  ores  and  in  the  ash  of  the  fuel.  Again, 
the  oxides  of  the  alkaline  earth  metals,  particularly  of  lime, 
are  always  present  in  the  slag,  since  CaO  is  used  as  a  slag- 
ging flux  (cf.  p.  181).  The  chief  constituent  of  blast  furnace 
slag,  therefore,  is  a  calcium-aluminium  silicate,  and  this 
silicate  is  approximately  neutral  or  weakly  basic,  according 
to  the  impurities  in  the  raw  materials  and  according  to 
the  desired  nature  of  the  iron  to  be  produced,  and  it 
ss  more  or  less  rich  in  other  metallic  oxides  (FeO,  MnO,  etc.). 


Fie.  179.  —  Gray  Cast  Iron.    Graphite  (Black);  Cementite  (White)  with 
Lamellar  Pearlite 


The  basic  slags  crumble  readily  after  exposure  to  the 
atmosphere,  but  when  it  is  desired  to  utilize  the  less  basic 
slags  they  are  usually  granulated  by  allowing  them  to 
flow  into  water  as  they  are  tapped  from  the  furnace.  The 
more  basic  slags,  particularly  of  foundry  iron,  when  they 
contain  suitable  amounts  of  SiC>2,  A12O3  and  CaO,  are 
used  in  the  preparation  of  Portland  cement. 
FURNACE  GAS.  In  discussing  the  reactions  of  the  blast 
furnace,  it  was  mentioned  that  a  gas  escapes  from  the 


IRON  195 

top  of  the  furnace  containing  up  to  28%  of  CO  (p.  187). 
One  cubic  foot  of  such  a  gas  possesses  a  heating  value 
of  89  B.T.U.  For  one  ton  of  pig  iron,  there  are  about 
160,000  cu.ft.  of  this  gas.  A  part,  half  at  the  most,  serves 
to  heat  the  blast  that  is  to  be  used  in  the  furnace,  and 
there  remains  from  80,000  to  90,000  cubic  feet  of  the  gas 
for  other  purposes.  Until  recently  this  gas  was  used  for 
heating  boilers  which  served  as  a  source  of  power  for 


FIG.  180. — Cast  Iron  showing  Temper  Carbon.     (Black);  in  Ferrite  (White), 
Surrounded  by  Lamellar  Pearlite. 

moving  the  raw  materials  and  the  products  of  the  fur- 
naces; to-day  the  gas  is  being  utilized  chiefly  in  gas  engines. 
The  blast-furnace  industry,  however,  requires  but  a  quar- 
ter of  the  power  that  can  be  produced  thus,  and  if  this 
additional  power  is  not  needed  for  other  work  in  the 
plant  (e.g.,  for  rolling  milJs)  it  can  serve  for  outside  pur- 
poses (e.g.,  for  electric  plants).  The  amount  of  power  thus 
available  can  be  easily  computed.  After  making  the  deduc- 
tions as  just  outlined  for  the  gas  that  is  utilized  in  connec- 


196  METALLURGY 

tion  with  the  blast-furnace  work,  there  remains  75%  of 
80,000  cubic  feet  of  gas,  or  from  a  3oo-ton  furnace, 
60,000X300=18,000,000  cubic  feet  of  gas  in  24 hours;  125 
cubic  feet  of  the  gas  yield  approximately  i  horse  power  hour. 
Thus  18,000,000  cubic  feet  from  a  3oo-ton  furnace  will  yield 

18,000,000 

hourly  about  —  — =  6000  horse  power  hours. 

125X24 

(C)   Iron  Alloys  and  Iron  Compounds 

Besides  pig  iron,  certain  alloys  of  iron  and  iron  compounds  to  be 

used  in  the  preparation  of  different  kinds  of  forgeable  iron  and  steel 

are  obtained  by  smelting  in  blast  furnaces. 

Spiegeleisen  and  Ferromanganese,  see  Manganese. 

Ferro-Chrome,  see  chromium. 

Ferro-Molybdenum  is  obtained  in  the  process  of  Borchers  and 
Lehmer  by  smelting  molybdenite  (MoS2)  in  an  electric  furnace 
with  lime,  carbon,  and  Fe  or  a  sulphide  or  oxide  of  iron. 

Ferro-Vanadium  is  prepared  by  a  reducing  fusion  of  ferrous 
vanadate  with  carbon  and  enough  Fe  in  the  flux  so  that  the 
resulting  alloy  will  contain  not  more  than  25%  Vd.  The  fer- 
rous vanadate  is  obtained  from  sodium  vanadate,  which  is 
prepared  by  an  oxidizing  roast  of  vanadium  ores  with  Na^COs 
or  NaCl,  or  by  fusing  lead  vanadate  with  soda;  the  sodium 
vanadate  is  extracted  with  water  and  the  clarified  solution  is 
treated  with  FeSO.i  solution. 

Ferro-Silicon  can  be  produced  like  pig  iron  in  the  blast  fur- 
nace by  smelting  iron  ores  with  a  gangue  or  flux  rich  in  SiO2. 
In  such  cases  the  aim  is  to  produce  a  slag  rich  in  AUOa.  By 
the  blast  furnace  method  the  Si  content  may  be  made  as  high 
as  about  16%.  Richer  alloys,  with  25%,  50%  or  more  of  Si,  are 
usually  obtained  in  electric  furnaces  by  the  reduction  of  mix- 
tures of  iron  ores  and  sand.  Ferro-silicon  has  also  been  obtained 
as  a  by-product  in  the  manufacture  of  CaC2.  In  this  case  a 
coal  with  an  ash  high  in  SiO2  was  used.  Iron  ore  was  then 
added  to  the  charge  and  a  ferro-silicon  with  about  20%  Si 
obtained. 


IRON  197 

(D)    Forgeable  Iron 

The  equilibrium  diagram,  Fig.  176,  p.  190,  contains  a  point 
a,  which  corresponds  to  2%  C,  and  represents  a  solid  solution  of 
the  composition  2Fe3C  +  i5Fe.  This  is  the  saturation  point  of 
f-iron  for  Fe3C  and  represents,  at  the  same  time,  the  upper  carbon 
limit  for  the  forgeability  of  iron.  Although  in  practice  the  upper 
limit  for  a  forgeable  iron  is  taken  as  1.6%  C,  still  in  the  light 
of  the  relations  between  Fe  and  C  which  have  been  sufficiently 
cleared  up  scientifically,  we  may  regard  as  forgeable  iron  all 
alloys  lying  between  pure  Fe  and  the  saturated  solution  of  2Fe3C 
in  i5Fe. 

At  S  the  C  content  amounts  to  0.9%,  corresponding  to  a 
composition  of  Fe3C  +  2oFe  (pearlite).  Although  with  a  C  content 
considerably  lower  than  0.9%,  there  is  an  increase  in  hardness 
when  the  metal  is  heated  and  then  rapidly  cooled,  this  increase 
in  hardness  by  such  treatment  begins  to  be  very  marked  at  0.9% 
C.  At  this  point  a  solution  is  reached  from  which,  on  slow 
cooling,  pure  Fe  (ferrite)  can  no  longer  crystallize  out  as  an 
independent  constituent  of  the  alloy,  but  is  contained  only  in  the 
eutectoid  called  pearlite,  Fe3C  +  2oFe;  with  higher  C  content  the 
eutectoid  is  formed  in  the  presence  of  Fe3C. 

The  characteristic  hardening  of  steel  has  been  explained 
by  assuming  a  transformation  of  pearlite  into  martensite  by 
heating  and  maintaining  this  solid  solution  by  quenching;  on 
the  other  hand,  the  transformation  of  steel  into  the  malleable  con- 
dition has  been  attributed  to  a  transformation  of  martensite 
into  pearlite,  either  by  reason  of  slow  cooling  or  by  keep- 
ing the  quenched  steel  at  moderate  temperatures  (below  700°) 
for  a  long  time.  The  more  the  C  content  falls  below  0.9%,  the 
more  ferrite  exists  in  the  presence  of  pearlite.  It  is,  however, 
easy  to  understand  why  a  steel  with  less  than  0.9%  C  can  be 
hardened.  When  0.9%  C  is  present,  the  sample  can  contain  100% 
of  pearlite,  and  therefore  a  steel  with  0.3%  C  contains  33.3%  of 
pearlite  which  can  be  transformed  into  martensite.  On  the 
other  hand,  it  is  plain  that  the  hardening  power  and  malleability 


198  METALLURGY 

will  not  disappear  when  the  C  content  is  above  0.9%,  for  beside 
the  cementite,  which  now  occurs  in  ever-increasing  amount, 
it  is  still  possible  for  pearlite  to  be  formed  and  it  is  not  until  2%  C 
is  reached  that  the  amount  of  pearlite  is  so  small  in  comparison 
with  the  amount  of  cementite  present  that  the  malleability  prac- 
tically ceases. 

The  property  of  hardening  can  be  caused  with  very  low  C 
content  by  the  presence  of  other  substances  such  as,  for  example, 
Cr  and  W. 

The  transformation  of  pig  iron  into  forgeable  iron  must  evidently 
consist  in  diminishing  the  amount  of  C  present  and  removing 
as  completely  as  possible  the  other  undesirable  constituents 
of  the  crude  iron.  In  all  cases  this  is  accomplished  by  oxidation 
processes.  In  one  case  (malleableizing)  only  C  is  removed,  but 
in  all  other  cases,  other  constituents  are  removed  simultaneously. 
At  the  same  time  it  is  impossible  to  prevent  a  part  of  the  Fe  itself 
from  being  oxidized  and  passing  partly  into  the  slag  and  partly 
into  the  purified  iron  as  FeO,  so  that  a  further  reduction  is 
necessary  to  remove  the  oxide. 

Malleableizing.  In  1903  to  1905  F.  Wlist  and  some  of  his 
students  proved  experimentally  that  by  exposing  white  iron  for 
a  long  time  to  a  temperature  of  about  1000°  (annealing)  there 
is  a  decomposition  of  the  dissolved  Fe3C  into3Fe  +  C  in  such 
a  way  that  the  C  is  not  deposited  as  graphite,  as  in  the 
formation  of  gray  iron,  but  entirely  in  an  amorphous  condition 
such  that  it  is  readily  susceptible  to  oxidation.  As  oxidizing 
agents,  oxide  iron  ores,  usually  Fe2Os,  but  sometimes  FeCO3, 
are  employed.  The  iron  objects  to  be  malleableized,  already 
in  the  form  of  finished  castings,  are  packed  in  this  ore, 
gradually  brought  to  a  temperature  lying  between  900°  and 
1000°  and  kept  there  about  a  week.  Until  recently  it  has 
been  held  in  metallurgical  literature  that  the  oxidation  of  the 
C  begins  at  the  surface  where  the  -oxidized  ore  is  directly  in 
contact  with  the  iron,  and  that  then  the  solid  C  migrates  from 
the  interior  to  the  carbon-free  outer  layers,  where  it  is  oxi- 
dized in  turn.  The  author,  in  his  Inorganic  Chemistry,  opposed 


IRON 


199 


this  view  in  1893;  in  1908  Wiist  submitted  indisputable 
experimental  proof  that  the  decarburization  is  brought  about 
by  the  aid  of  CO2,  which  is  obtained  partly  by  the  contact  of 
particles  of  iron  containing  C  with  the  Fe2O3,  and  partly  from 
the  combustion  gases  of  the  furnace.  Wiist  showed,  more- 
over, that  these  gases  diffused  into  the  interior  of  the  castings, 
where  the  most  favorable  conditions  exist  (cf.  diagram  174) 
for  the  progress  of  the  reaction  CO2  + C  =  2CO,  while  the  CO 
thus  formed  is  again  converted  into  CO2  by  the  oxidized  ore, 
and  can  once  more  take  part  in  the  reaction. 


FIG.  181. — Charcoal  Hearth. 

Production  of  Wrought  Iron.     These  methods  of  refining  iron 
are  the  oldest  in  the  development  of  iron  metallurgy. 
Among  them  are: 

The  Charcoal  Hearth  Process.  This  is  still  used  in  some 
places  (Sweden,  Styria,  Russia)  where  a  sufficient  supply 
of  charcoal  can  be  obtained  cheaply.  The  "  hearths  "  or 
"  forges "  contain  a  hollow  pit,  usually  lined  with  iron 
plates,  and  are  provided  with  a  tuyere  in  the  high  back  wall. 
The  cast  iron  is  melted  upon  charcoal  but  with  a  blast  of 
air  sufficient  for  the  necessary  oxidation,  and  directed  so 
that  the  liquid  iron  must  drop  through  the  blast  zone.  A 


200 


METALLURGY 


glance  at  the  diagram,  Fig.  176,  shows  that  the  melting 
^  pointjofiron  rises  as  its  purity  increases.  The  iron,  that 
was  at  first  liquid,  collects  in  the  bottom  Tfearth  as  a  pure 
viscous  metal  (blooms)  which  is  freed  from  any  included 
slag  by  mechanical  working  (hammering,  rolling),  and 
welded  to  a  compact  mass. 

Puddling.  This  is  essentially  the  same  kind  of  work,  but 
differs  by  melting  the  cast-iron  in  a  reverberatqry^jurnace, 
in  which  the  oxidation  is  assisted  by  the  presence  of  iron 
oxides,  added  in  the  form  of  slags  rich  in  Fe3O4,  and 
waste  from  the  mechanical  working  of  forgeable  iron  (e.g., 


FIG.  182. — Longitudinal  Section  of  Puddling  Furnace. 

hammer  scale).  The  contact  of  these  oxides,  and  of  the  O 
from  the  flame  gases,  is  aided  by  stirring  up  the  slag  with 
rabbles.  As  the  pure  iron  solidifies  it  is  broken  up  from 
time  to  time  and  turned  over,  etc.,  so  as  to  expose  the  whole 
charge  to  the  refining  action.  As  in  the  last-mentioned  proc- 
ess, the  metal  is  finally  freed  from  slag  by  hammering,  and 
is  thereby  welded  together  into  a  conglomerate  of  Fe  crystals. 
The  reverberatory  furnaces  consist  of  beds  or  hearths  sup- 
ported by  iron  plates,  with  a  free  space  underneath  so  that 
they  are  exposed  to  the  cooling  action  of  the  atmosphere. 
The  side  walls  are  hollow  iron  plates,  which  can  be  cooled 
with  water,  and  are  lined  with  a  basic  slag  which  has  a 


IRON 


201 


high  melting  point,  and  is  rich  in  FeO.  The  older,  single- 
hearth  reverberatory  furnaces  have  a  large  grate  area  and 
the  waste  gases  which  pass  off  at  a  temperature  of  1200  to 
1300°  C.  are  used,  as  a  rule,  for  making  steam.  The 
newer  furnaces  have  regenerative  gas  firing  and  double 
hearths.  According  to  the  requirements  placed  upon  the 
iron,  10  to  20  charges  of  500  to  700  Ibs.  each  are  worked 
up  in  a  day  in  the  older  furnaces;  in  the  newer  ones,  24 
charges  of  1000  Ibs.  each.  Loss  of  metal,  6  to  15%. 
For  heating  the  furnaces  and  for  power  sufficient  to  hammer 


FIG.  183. — Plan  of  Puddling  Furnace  with  Flat  Grate,  burning  Anthracite  Coal. 

or  roll  the  blooms,  the  coal  consumption  equals  the  weight 
of  the  finished  product. 

The  forgeable  iron  produced  by  either  of  the  last  two  processes 
is  called  wrought  iron.  Since  the  temperature  of  the 
furnace  hardly  reaches  1300°,  it  is  evident  that  the  crystals 
of  iron  which  are  formed  during  the  refining  cannot  fuse 
together  but  become  welded. 

Oxidizing  Fusion  of  Cast  Iron  Above  the  Melting-point  of 
Pure  Iron.  Production  of  Steel.  The  oxidation  is  accom- 
plished by  atmospheric  oxygen,  as  well  as  by  the  oxygen  of  iron 
oxides,  by  blowing  in  converters;  or  by  smelting  in  regener- 
ative furnaces,  or  in  electric  furnaces. 
The  blowing  consists  in  conducting  a  blast  of  air  through  the 


202  METALLURGY 

molten  cast  iron  until  the  impurities  have  been  burned  away; 
the  heat  of  combustion  serves  to  maintain  a  temperature 
so  high  that  the  purified  iron  remains  liquid.  The  substances 
present  have  the  following  heats  of  combustion  per  pound: 
81=14,090  B.T.U.;  P=  10,740  B.T.U.;  0  =  4297  B.T.U.; 
Mn  =  3io3  B.T.U.,  and  Fe  =  2435  B.T.U.  The  rise  in 
temperature  caused  by  these  elements  is  in  the  following 
order:  By  the  combustion  of  i%  of  the  entire  weight  of  the 
charge,  Si  increases  the  temperature,  according  to  recent 
calculations  of  Wiist,  287°  C.  ;  the  same  amount  of  P,  185°  C.  ; 
of  Mn,  61°  C.;  of  Fe,  44°  C.;  and  of  C.,  8.8°  C.  (CO 
escapes  as  a  gas  and  takes  along  with  it  the  greater 
part  of  the  heat  of  combustion).  Therefore,  for  purification 
by  this  process,  an  iron  rich  in  Si  or  in  P  is  desirable.  An 
iron  containing  large  amounts  of  both  these  elements  could 
not  be  refined  satisfactorily  in  a  converter,  because  a  cast  iron 
rich  in  Si  yields  an  acid  slag  and  necessitates  an  acid  lining 
in  the  melting  apparatus.  The  SiO2  formed  by  the  combus- 
tion, and  that  present  in  the  converter  lining,  will  tend  to 
prevent  the  slagging  off  of  P2O5,  for  it  would  set  P2C>5  free 
from  any  Fe3(PO4)2,  in  accordance  with  the  equation: 

Fe3(P04)2  +  3Si02  =  3FeSiO3  +  P2O5. 


Free  P2Oo,  however,  is  readily  reduced  by  the  large  excess 
of  hot  iron  present  and  the  P  will  thus  pass  into  the  iron 
again  in  the  form  of  phosphide.  The  process,  therefore, 
is  conducted  according  to  one  of  two  principles,  the  older 
one  being  named,  after  its  discoverer,  the 

Bessemer  Process.  In  this  case  pig  irons  high  in  Si  (up  to 
2%  Si)  are  "  blown  "  in  tilting,  pear-shaped  vessels  called 
Bessemer  converters.  These  are  made  of  wrought-iron 
plates  bolted  firmly  together  and  are  lined  with  an  infusible 
silicious  rock  termed  ganister.  They  are  fitted  with  tuyere 
boxes  at  the  bottom.  In  the 

Basic  Bessemer  or  Thomas-Gilchrist  Process,  a  cast  iron  rich 
in  P  is  used  (up  to  2.5%  P)  with  as  low  Si  as  possible.  The 


IRON  203 

converters,  in  this  case,  are  given  a  basic  lining  consisting 

of  calcined  dolomite  with  the  addition  of  lime.     The  slags 

formed  consist  of  a  basic  calcium  phosphate,   with  P2O5 

up  to  25%;   this  forms  a  valuable  by-product,  the  so-called 

Thomas  slag,  which  is  used  as  a  fertilizer. 

The  converters  are  built  in  different  sizes,  up  to  10  ft.  in 

height  and  10  ft.  in  width  (measuring  the  shell  without 

lining).     They  will  hold  from  1.5  to  25  tons  of  pig  iron, 

which  is  blown  in  from  15  to  20  minutes.     The  lost  metal 

amounts  to  iotoi5%.     As  adjunct  to  the  Bessemer  plant, 

there  remain  to  be  mentioned  the 

PiG-lRON  MIXER,  which  is  a  revolving  drum  on  a  horizontal 

axis;    it  is  made  of  iron  plates  with  a  silicious  lining,  and 

has  a  capacity  up  to  600  tons.     By  this  mixing,  a  fairly 

uniform  material  is  guaranteed  for  the  converter  and  a 

certain  preliminary  purification  takes  place,  inasmuch  as 

Mn  combines  with  S  and  the  MnS  separates  out  as  slag. 

Open  Hearth  Process.     This  process  was  rendered  possible  by 

the  discovery  of  regenerative  firing  by  Friedrich  and  Wilhelm 

Siemens  and  first  utilized  by  Emil  and  Pierce  Martin  in  a 

Siemens  furnace  for  the  production  of  steel  (1865). 

In  the  Siemens-Martin   Process  the  oxidation  of  the  impurities 

in   the   iron   is   effected   partly  by   the   oxygen    of    the    air 

and  partly  by  the  oxides  added,  which  is  either  rust  on  scrap 

iron,  or  in  the  form  of  hammer-scale,  or  pure  ore  (magnetite, 

hematite).     According  to  whether  the  raw  material  is  rich 

in  Si  or  in  P,  the  work  is  either  carried  on  "  acid  "  or  "  basic," 

as  in  the  Bessemer  process,  and  the  lining  of  the  hearth 

is  governed  in  the  same  way. 

While  the  Siemens-Martin  process  was  formerly  used 
almost  exclusively  for  working  up  pig  iron  or  wrought 
iron  into  steel,  it  has  since  been  used  with  much  success  for 
smelting  pig  iron  with  iron  ores;  or,  in  other  words,  a  method 
of  working  has  developed  in  which  considerable  amounts 
of  steel  are  obtained  directly  from  ores.  Such  processes 
are  the  recent  ones  of  Bertrand-Thiel  and  of  Talbot.  In 


204 


METALLURGY 


the  former,  the  work  is  carried  out  with  two  Siemens  furnaces 
with  basic  hearths  in  such  a  way  that  in  one  furnace  a 
preliminary  refining  takes  place  with  the  production  of  a  slag 


FIG.  1840. — Converter. 

rich  in  P2O5,  and  in  the  second  the  finishing  work  is  carried 
out  with  a  fresh  addition  of  ore.  The  Talbot  process  works 
with  one  large  furnace  (up  to  250  tons  charge)  in  which  the 


IRON 


205 


liquid  pig  iron  is  run  upon  the  previously-introduced  ore  and 
lime.  After  the  conclusion  of  the  violent  reaction  that  takes 
place,  the  first  slag  is  removed  in  order  to  smelt  the  metal 
further  with  another  addition  of  ore. 


FIG.  1846. — Converter. 

APPARATUS:  Reverberatory  furnaces  with  open  hearths  and 
with  regenerative  firing.  The  hearth  rests  upon  iron 
plates  and  consists  either  of  lump  quartz  or  of  calcined 
dolomite.  The  other  walls  of  the  smelting  space  are  also 
of  acid  or  basic  material,  at  least  on  the  inside.  The 
side  walls  contain  apertures  for  working,  charging,  and 
tapping;  the  end  walls  have  openings  for  admitting  the 


206 


METALLURGY 


gas  and  air  which  are  alternately  introduced,  from  one 
side  and  then  the  other,  after  passing  through  the  heating 


FIG.  185. — Siemens  Open  Hearth  Furnace. 


FIG.  186. — Siemens  Open  Hearth  Furnace. 

chambers  lying  beneath  the  hearth.  Each  furnace,  therefore, 
has  two  pairs  of  heating  chambers  which  consist  of  shaft- 
like  structures  latticed  with  firebricks  to  aid  in  the  trans- 


IRON  207 

ference  of  heat.  One  chamber  of  each  pair  serves  for 
heating  the  combustible  gases,  and  the  other  for  heating 
the  air  required  in  the  combustion  of  these  gases.  During 
the  operation  of  the  furnace,  therefore,  the  current  of  gas 
is  sent  into  one  chamber  from  the  bottom,  and  the  air  is 
sent  into  the  other  chamber.  Both  of  these  currents  pass 
through  the  narrow  openings  in  the  brickwork  and  are 
heated  by  contact  with  the  hot  bricks.  The  two  streams 
are  kept  separate  until  after  they  leave  the  top  of  the 
chambers,  when  they  are  introduced  into  the  smelting 
space  and  combustion  at  once  takes  place.  Since  a  high 
temperature  is  required  in  the  smelting  zone,  the  heat 
transference  cannot  be  carried  very  far  here.  Hence,  the 
hot  waste  gases,  as  they  leave  this  zone,  are  led 
through  channels  in  the  opposite  wall,  into  the  second 
pair  of  heating  chambers  in  which,  as  they  pass  down- 
ward and  out  at  the  bottom,  they  heat  the  bricks  and 
walls.  After  an  interval  of  30  to  50  minutes  the  direc- 
tion of  the  gas  currents  is  reversed  and  the  air  and  com- 
bustible gases  are  now  conducted  through  the  chambers 
that  have  just  been  heated  by  the  escaping  gases. 

The  furnaces  are  sometimes  made  so  that  they  will  tilt 
and  are  then  suited  for  very  large  charges.  The  ordinary 
furnaces  with  fixed  hearths  have  a  capacity  of  from  12  to 
40  tons,  but  the  tilting  furnaces  will  hold  up  to  300  tons. 
From  three  to  six  charges  are  worked  through  in  24  hours. 
Metal  losses,  6  to  8%. 

Bessemer  and  Open  Hearth  Treatment  in  Succession  (Duplex 
Process)  has  been  practiced  in  some  places  (e.g.,  in  Wittko- 
witz)  for  working  up  a  pig  iron  containing  too  much  P  to  be 
satisfactorily  purified  by  the  ordinary  Bessemer  process,  and 
too  much  Si  for  the  basic  Bessemer  process. 
Electric  Smelting,  like  the  Siemens-Martin  process,  can  be  used 
for  working  up  either  scrap  or  pig  iron.  In  1878,  soon  after 
the  discovery  of  the  dynamo,  Wilhelm  Siemens  proposed  to 
smelt  iron  by  electricity,  and  since  that  time  there  has  been 


208  METALLURGY 

no  end  ot  experiments  toward  solving  the  problem,  but  up 
to  1900  these  were  mostly  concerned  with  the  question  of 
electrodes.  In  heating  by  the  electric  arc,  carbon  electrodes 
are  indispensable.  When,  however,  the  electric  arc  springs 
from  a  carbon  electrode  to  the  mass  to  be  heated,  it  is  un- 
avoidable, if  this  mass  consists  of  a  metal  like  iron,  that 
some  carbon  will  be  taken  up  by  it.  The  taking  up  of 
carbon,  in  spite  of  the  use  of  two  carbon  electrodes,  was  first 
prevented  satisfactorily  by  means  of  the 


FIG.  187. — Heroult  Furnace. 

HEROULT  FURNACE,  consisting  of  an  electric  arc  and  resist- 
ance furnace  in  which  both  electrodes  are  introduced  from 
above  into  the  smelting  hearth.  An  oxidizing  slag  (mag- 
netite and  a  basic  flux)  is  kept  upon  the  metal.  The 
electrodes  are  placed  far  enough  apart,  and  the  layer  of 
slag  on  the  other  hand  is  kept  so  thin  that  the  current 
passes  from  one  electrode,  with  the  formation  of  an  arc, 
first  to  the  slag,  then  to  the  metal,  from  the  other  end  of 
the  metal  again  to  the  slag,  and  thence  to  the  other  elec- 


IRON  209 

trode  with  the  formation  of  a  second  arc.  Evidently, 
then,  as  sources  of  heat  there  are:  two  electric  arcs  playing 
directly  upon  the  surface  of  the  slag,  two  layers  of  slag 
resting  directly  upon  the  metal,  and  the  metal  itself  as 
heating  resistance.  The  slag,  containing  the  substances 
capable  of  entering  into  chemical  reaction,  is  heated  very 
hot  and  so  is  the  metal  itself,  which  is  well  protected  from 
loss  of  heat.  The  C  from  the  electrodes  is  oxidized  in  the 
slag  layer.  A  glance  at  the  whole  arrangement  shows,  and 
this  has  been  confirmed  in  practice,  that  the  oxidizing 
agent  in  the  refining  slag  is  brought  by  the  intense  heat 
to  its  maximum  reaction  velocity  in  respect  to  the  impurities 
present  in  the  bath  of  iron.  As  a  matter  of  fact,  it  is 
possible  to  accomplish  in  this  furnace,  and  in  that  of  the 
following  ones,  a  much  more  efficient  purification  of  the 
iron  than  in  any  of  the  previously  described  furnaces.  The 
potential  required  for  the  Heroult  furnace  is  no  to  120 
volts. 

GIROD  FURNACE.  This  arc-resistance  furnace  was  designed 
by  Paul  Girod-Ugine  and,  as  compared  with  the  Heroult 
furnace,  excels  in  simplicity  of  construction  and  operation. 
One  pole  of  the  electric  arc,  consisting  of  one  or  more 
blocks  of  carbon,  is  introduced  from  above  into  the  middle 
of  the  smelting  hearth.  The  slag  floating  upon  the  metal 
forms  the  other  pole.  It  obtains  its  contact  through  the 
metal  bath,  which  is  placed  in  circuit  with  iron  rods,  or 
rings,  that  are  introduced  from  below  into  the  furnace  near 
the  periphery  of  the  metal  bath.  These  connecting  pieces, 
which  are  cooled  somewhat  outside  the  furnace,  are  so 
dimensioned  that  they  can  just  conduct  the  current  without 
too  great  resistance;  in  other  words,  they  are  given  as 
much  load  as  possible  so  that  there  is  a  good  distribution 
of  the  current  all  over  the  bath.  The  ends,  drawn  out 
somewhat  where  they  are  in  contact  with  the  bath,  are 
kept  just  hot  enough  by  the  current  to  prevent  any  undesired 
conducting  away  of  heat  from  the  bath.  As  sources  of 


210 


METALLURGY 


FIG.  188. — Girod  Furnace. 


IRON  211 

heat  in  the  Girod  furnace,  therefore,  there  are:  one  or 
more  electric  arcs  over  the  middle  of  the  bath,  a  slag  layer 
upon  the  metal,  and  the  metal  itself.  The  current 
passes  with  the  formation  of  an  arc  from  one  or  several 
upper  electrodes  through  the  refining  slag  to  the  middle 
of  the  metal,  whence  it  radiates  uniformly  in  all  directions, 
heats  the  bath,  and  keeps  it  in  motion,  and  is  carried 
away  in  the  outer  periphery.  For  the  Girod  furnace,  a 
potential  of  55  to  65  volts  is  necessary. 

INDUCTION  FURNACES,  a  discovery  of  de  Ferrantis  in  the 
year  1887,  were  first  adapted  by  Kjellin,  and  almost  at 
the  same  time  by  Colby  (1900),  for  use  as  furnaces  for 
iron  smelting.  A  closed  electro-magnet,  built  in  a  right 
angle,  forms  a  transformer  with  the  fused,  or  ready- to-be- 
fused,  iron  which  rests  in  a  ring-shaped  gutter  that  is 
placed  perpendicularly  to  the  iron  core  of  the  magnet.  The 
most  effective  construction  has  proved  to  be  that  in  which 
the  winding  of  the  electro-magnet  lies  concentric  to  the 
iron-bath,  whether  this  be  within  or  without  the  field  of 
the  primary  circuit.  The  secondary  current  produced 
in  the  iron  of  the  gutter  by  an  alternating  current  in  the 
primary  circuit  is  at  once  transformed  into  heat  and  the 
iron  is  quickly  melted.  Since  the  furnace  works  without 
electrodes,  there  is  no  danger  of  contamination  from 
electrode  material,  but  notwithstanding  this,  the  majority 
of  these  induction  furnaces  have  not  proved  satisfactory. 
The  oxidizing  slag  (iron  oxides,  etc.)  owing  to  its  low 
conductivity,  takes  but  little  part  in  the  formation  of  the 
secondary  current  and  in  the  resulting  heat  transference, 
but  obtains  its  heat  from  the  iron.  The  slag,  therefore, 
is  considerably  colder  than  that  of  the  Heroult  and  Girod 
furnaces.  In  those  furnaces,  the  slag  acts  as  one  pole  of 
the  arc  and,  because  of  its  poor  conductivity,  causes  the 
production  of  much  heat  in  the  passage  of  electricity 
through  it,  and  this  high  temperature  is  particularly 
favorable  for  increasing  the  velocity  of  the  reactions  that 


212 


METALLURGY 


FIG.  189. — Rochling-Rodenhauser  Furnace. 

n 


FIG.  190. — Rochling-Rodenhauser  Furnace. 


IRON 


213 


tend  to  take  place  between  the  oxygen  of  the  slag  and  the 
impurities  of  the  iron.  This  difficulty  is  overcome  in  the 
furnace  of  ROCHLING  and  RODENHAUSER  by  giving  direct 
resistance  heating  to  a  part  of  the  bath.  Both  vertical 
arms  of  the  magnet's  iron  core  H  (Fig.  191),  are  provided 
with  windings,  A,  of  the  primary  circuit,  and  with  melting 
gutters  C;  these  last  unite  in  a  broad  middle  hearth,  D. 
At  the  ends  of  this  middle  hearth  are  two  protective  walls 


FIG.  191. — Rochling-Rodenhauser  Furnace. 

of  magnesia  or  dolomite,  both  of  which  materials  conduct 
well  at  high  temperatures.  Back  of  these  walls  and  con- 
nected with  a  secondary  current  J5,  are  pole  plates  E  from 
which  currents  are  said  to  pass  through  the  charge  con- 
tained in  the  broad  hearth. 

CAPACITY  OF  ELECTRIC  FURNACES.  Whereas  the  Heroult 
and  Girod  furnaces  work  advantageously  even  with  cold 
material,  the  induction  furnaces  are  practically  restricted 


to  work  with  molten  and  partly  purified  iron. 


Starting 


214  METALLURGY 

with  cold  metal  as  raw  material,  in  a  small  Heroult  or 
Girod  furnace  (2  to  2.5  tons  charge)  about  900  kilowatt 
hours  =  1230  horse-power  hours  are  to  be  reckoned  per  ton 
of  finished  steel;  in  large  furnaces  (8  to  10  tons),  700  kilo- 
watt hours  =  95 1  horse-power  hours.  In  the  first  case,  there- 
fore, about  0.14,  and  hi  the  second  0.108  horse-power  year 
per  ton  of  steel.  If  the  furnace  can  be  charged  with  liquid 
metal,  then  the  expenditure  of  energy  is  diminished  by  at  least 
one-third.  The  Girod  method  of  introducing  the  electric 
current  gives  greater  velocity  and  uniformity  when  working 
with  cold  metal  than  does  the  Heroult  method.  The 
consumption  of  power  by  the  induction  furnace  is  less 
only  when  the  iron  is  already  melted  and  partly  purified; 
when  cold  metal  is  used,  more  power  is  required  than 
with  the  Heroult  and  Girod  furnaces. 

In  all  processes  for  the  production  of  steel,  it  is  possible 
to  remove  all  the  impurities  only  by  an  over-oxidation,  so  to 
speak,  i.e.,  as  the  concentration  of  the  impurities  diminishes, 
the  more  Fe  will  be  oxidized  to  FeO,  which  partly  passes  into 
the  slag  (metal  loss)  and  is  partly  held  in  solution  by  the  iron. 
If  to  this  iron,  containing  FeO,  the  various  substances  C,  Si, 
etc.,  were  added  which  are  necessary  for  the  production  of 
special  kinds  of  steel,  then,  particularly  with  C,  chemical 
reaction  would  take  place  during  the  solidification,  and  in  the 
case  of  C  this  reaction  would  be  accompanied  by  the  evolu- 
tion of  a  gas  (CO)  and  the  result  would  be  a  porous  metal. 
FeO,  remaining  dissolved  in  the  iron,  makes  the  metal  red-short; 
(i.e.,  brittle  when  hot).  After  the  fusion  is  completed,  there- 
fore, there  is  in  all  cases  a 

Reduction  by  the  addition  of  Mn  in  the  form  of  a  Mn-Fe 
alloy,  of  C  for  the  recarburization,  of  Si  in  the  form  of  silicides, 
and  finally  of  Al  for  the  removal  of  the  last  traces  of  FeO 
and  for  accelerating  the  reactions  between  FeO  and  the 
other  substances  charged.  The  addition  of  the  Al  takes  place 
usually  in  the  ladles,  or  even  after  pouring  into  the 
molds, 


IRON  215 

The  Finishing  of  Forgeable  Irons.     In  the  refining  methods 
which  have  been  described,  it  is  practically  impossible  to  stop 
the  process  when  the  metal  contains  the  desired  amount  of  C 
or  Si  corresponding  to  a  certain  quality  of  iron.     The  impurities, 
are,  therefore,  removed  completely  and   then   substances   are 
added  which  will   impart    to  the  iron  the  desired  properties. 
The  simplest  and  quickest  way  to  do  this,  is  by 
Alloying    in   the    Bath,   which  is   accomplished   in   the   pro- 
duction of  steel  by  adding  the  necessary  substances   them- 
selves to  the  bath  of  pure   Fe,  or  in  the  form  of  an  alloy 
or  chemical  compound,  either  in   the  smelting  furnace  or 
in  the  ladle.     For  the  recarburization  of  iron  up  to  a  certain 
definite  point,  it  is  sufficient  to  add  pure  coke  or  graphite 
to  the  ladle  as  the  iron  is  poured  from  the  converter  (Darby- 
Phoenix  Process).     It  is  more  usual,  however,  to  add  spiegel- 
eisen  (see  Mn)  with  high  C  content.    Similarly,  for  introducing 
Si,   a    ferro-silicon    is   added.     Also   for   introducing   other 
metals,  such  as  Mn,  Cr,  W,  etc.,  it  is  customary  to  add  these 
in  the  form  of  an  iron  alloy  on  account  of  its  being  easier  to 
prepare  such  an  alloy   of   sufficient   purity  than  the  pure 
metal  itself;   W  and  Cr  as  vrell  as    Ni   and    other  metals, 
are  also  added  in  the  pure  state. 

For  producing  varieties  of  steel  of  very  uniform  compo- 
sition, particularly  when  it  is  not  necessary  to  keep  the  C 
content  low,  the  pure,  analyzed  metal  is  melted  with  the 
required  additions  (Ni,  W,  Cr,  etc.)  in  graphite  crucibles. 
In  this  way  crucible  steel  is  prepared,  and  recently  the  melting 
has  been  done  in  an  electric  furnace. 

By  Welding,  accomplished  by  pressing  together  (hammering 
or  rolling)  rods  of  different  kinds  of  iron  at  a  welding  heat, 
it  is  possible  to  prepare  different  qualities  of  iron  and  steel 
which  apparently  are  uniform  but  when  examined  under 
the  microscope  show  that  different  materials  have  been 
united  mechanically.  Particularly  for  the  so-called  dou- 
ble shear  steel  (for  cutlery,  etc.)  it  is  important  to  have 
hard  particles  embedded  in  a  softer  and  sufficiently  elastic 


216  METALLURGY 

matrix;   in  this  way  the  sharpness  of  instruments  is  main- 
tained. 

Cementation.  This  process,  like  that  of  malleableizing,  retains 
the  form  of  the  objects  subjected  to  it,  but  the  nature  of  the 
process  is  exactly  the  reverse.  The  objects  are  embedded 
in  charcoal  powder  and  heated  for  a  long  time  to  a  tempera- 
ture of  about  900°  C. ;  in  this  way  the  C  content  can  be  raised 
to  about  1.2%.  The  working  conditions  are  favorable 
for  the  reversal  of  the  reaction  that  takes  place  during  mal- 
leableizing; 2CO  =  C  +  CO2.  The  CO2,  as  fast  as  it  is 
formed,  is  reduced  by  the  bed  of  C  to  CO  again  and  can 
serve  anew  for  the  introduction  of  C  into  the  iron.  It  is 
also  possible  to  carburize  iron  objects  superficially  by  dip- 
ping them  in  baths  or  exposing  them  to  the  action  of  certain 
gases.  These  compounds,  or  gases,  on  being  heated,  give 
rise  to  C  or  carbides.  The  process,  called  case  hardening, 
is  accomplished  by  heating  the  metal  and  plunging  it  into 
some  substance  like  yellow  prussiate  of  potash  (potassium 
ferrocyanide)  K4Fe(CN)e  =  4KCN  +  FeC2  +  N2,  or  by  heat- 
ing in  a  current  of  illuminating  gas. 
Properties  of  Iron : 

SPECIFIC  GRAVITY:   7.86. 

COLOR:    grayish- white  with  high  luster. 

MECHANICAL  PROPERTIES:  tough,  very  ductile. 

STRUCTURE:  see  ferrite,  page  192,  Fig.  177. 

MELTING-POINT:   1512°  C.  (2754°  F.) 

BOILING-POINT:   2600°  C.  (?)  (4712°  F.) 

ELECTRICAL  CONDUCTIVITY:  about  0.14  that  of  Ag. 

MAGNETIC  PROPERTIES:    Fe  is  the  most  paramagnetic  of 

metais. 

ALLOYS  readily  with  most  of  the  earth  metals,  slightly  with 
Pb  and  Cu.  In  the  presence  of  Si,  Fe  will  take  up  more 
Cu;  or,  in  other  words,  it  dissolves  copper  silicide  more 
readily  than  it  does  pure  Cu.  Fe  also  alloys  easily  with 
its  own  compounds  with  metalloids  (C,  Si,  P,  S,  O)  (Cf. 
pig-iron,  page  188). 


IRON  217 

CHEMICAL  BEHAVIOR:  in  dry  air  it  is  fairly  stable  toward 
O  at  low  temperatures.  At  300°  C.  it  is  more  readily  oxi- 
dizable.  At  a  moderate  red  heat  it  combines  with  S  and  P, 
and  at  higher  temperatures  with  C  and  Si.  The  halogens 
act  energetically  upon  Fe  even  at  ordinary  temperatures. 
On  account  of  its  high  electrolytic  solution  tension  (+0.344 
toward  H)  it  dissolves  easily  in  dilute  mineral  acids 
with  evolution  of  H,  and  in  HNOs  with  reduction  of  the 
acid  and  evolution  of  nitric  oxide.  In  concentrated 
H2SO4  or  HNOs,  it  is  so  difficultly  soluble  that  iron  ves- 
sels can  be  used  for  transporting  these  acids.  In  the  tech- 
nically-important iron  compounds,  the  Fe  is  either  bivalent 
(ferrous  compounds)  or  trivalent  (ferric  compounds). 
Only  in  the  carbides  and  silicides,  which  are  formed  and 
are  stable  at  high  temperatures,  does  Fe  show  a  different 
valence.  The  most  important  carbide  has  the  compo- 
sition Fe3C  (cementite),  and  silicides  are  known  corre- 
sponding to  the  formulas  Fe2Si,  FeSi,  Fe3Si2,  and  FeSi2. 
The  properties  of  the  technically-important  varieties 
of  iron  have  already  been  discussed  sufficiently  in  the 
sections  denning  the  terms  cast  iron  (white  and  gray), 
wrought  iron,  and  steel. 


CHROMIUM 

Sources 

Natural  Sources : 

CHROME-IRON    ORE,   or   chromite,   FeCr2O4 

usually  accompanied  by   serpentine.     This   is   the   most 

important   chromium   ore.      Of   the   other   minerals,   the 

best-known   is 
CROCOITE,  lead  chromate,  PbCrO4,  now  scarcely  used  at  all 

for  the  production  of  metallic  chromium. 

Ferro-Chrome 

A  Reducing  Smelt  of  the  chrome -iron  ore  gives  directly  a  com- 
mercial Fe-Cr  alloy.  As  flux,  only  charcoal  or  coke  powder 
is  required.  At  the  beginning  of  the  reaction,  which  takes 
place  with  evolution  of  CO,  there  is  danger  of  the  mixed 
ore  and  carbon  falling  apart  on  account  of  the  difference  in 
specific  gravities;  to  prevent  this,  a  little  colophonium  or  pitch 
is  added  to  the  charge. 

On  being  heated,  the  pitch  glues  together  the  ore  and  carbon, 
forming  a  coke  at  a  high  temperature,  and  then  finally  the 
whole  mass  unites  to  one  large  lump.  The  charge  is  regulated, 
of  course,  by  the  nature  of  the  ore.  To  one  ton  of  ore,  from 
250  to  330  Ibs.  of  charcoal  powder  and  125  to  155  Ibs.  of 
powdered  colophonium  or  pitch  is  added. 
APPARATUS:  Crucibles  in  reverberatory  furnaces,  using  forced 

draft  or  regenerative  firing,  or  in  electric  furnaces. 
THE  CRUCIBLES  are  made  of  graphite  or  of  clay.    When  many 

of  the  former  material  are  required,  the  cost  of  operating  is 

materially  increased.     Clay  crucibles  can  be  used  but  once, 

218 


CHROMIUM 


219 


but  they  are  cheap.  In  the  simple  furnaces  the  con- 
sumption of  fuel  is  very  great  and  it  is  hard  to  reach  the 
necessary  temperature.  In  this  respect  the  regenerative 
furnaces  are  more  advantageous.  Since  in  using  clay  crucibles 
the  heating  chambers  must  ^be  cooled  enough  so  that 
there  is  no  danger  of  breaking  the  crucibles  in  recharg- 
ing the  furnace,  the  author  usually  combines  two  furnaces 


FIG.  192.— Section  A,  B,  C,  D,  E,  F. 


FlG   I93.— Section  G,  H,  K,  L5  M.          FIG.  194.     Section  N,  O,  P,  R. 
FIGS.  192-4. — Crucible  Furnace  with  Regenerative  Chambers. 

in  such  a  way  that  it  is  possible  to  work  alternately,  the 
furnace  with  the  fresh  charge  receiving  its  heat  from  the 
slowly-cooling  furnace  with  the  finished  charge.  The  furnace 
block  contains  the  gas  producer  and  the  two  regenerative, 
reverberatory  furnaces.  (Figs.  192  to  194.)  The  producer 
is  a  simple  shaft  furnace  with  horizontal  grate  for  coke. 
It  sends  its  gas  into  one  of  the  main  flues  of  the  adjacent 
reverberatory  furnace.  From  the  main  flue,  the  gas  can  be 


220  METALLURGY 

led  into  one  or  the  other  branching  flues  of  each  furnace 
by  uniting  two  faucet  tubes  with  a  n -shaped  tube  of  sheet 
metal.  The  regenerators  are  there  only  for  preheating  the 
air.  The  work  is  earned  out  as  usual,  changing  the  direction 
of  the  gas  every  half  hour  or  hour.  The  change  in  direction 
of  the  producer  gas  is  effected  by  the  above-mentioned  faucet 
tubes,  of  the  air  and  waste-gases  by  valves. 

Pure  Chromium 

To  prepare  Cr  comparatively  pure  and  free  from  Fe, 
I.  The  Iron  Must  be  Separated  Chemically  from  chromium. 
This  is  accomplished  by  the  following  operations: 

1.  Oxidizing  Roast  with  Alkaline  Fluxes.     Purpose:   changing 
FeO  to  Fe2Os,  Cr2O3  into  CrOs  and  union  of  the  latter  with 
CaO  or  Na2O. 

2FeCr2O4  +  7  O  +  4Na2CO3  =  Fe2O3  +  4Na2CrO4 + 4CO2. 

Contrary  to  the  views  prevailing  in  chemical  and  metallurgical 
literature,  the  mass  should  not  melt,  for  when  melting  takes 
place  the  surface  of  attack  for  the  O  upon  the  FeCr2O4 
is  greatly  lessened.  The  firing,  therefore,  does  not  need  to 
give  a  very  high  temperature.  Simple  reverberatory  furnaces 
(Figs.  195  to  197)  such  as  are  used  in  the  LeBlanc  process 
for  the  manufacture  of  soda,  are  suitable.  To  prevent  the 
charge  fusing  together,  CaCOs  has  been  added  or  used  to 
replace  a  part  of  the  Na2CC>3. 

2.  Lixiviation.     Purpose:     separation    of    soluble    chromate, 
Na2CrO4,    from    insoluble    Fe2Os    and    gangue.     As    sol- 
vent, hot  water  is  used  and  if  CaCrO4  is  present,  Na2CO3 
or  Na2SO4  is  added.     In  this  case,  the  work  is  performed 
in  closed  iron  drums  at  120°  to  130°  C.     The  filtered  Na2CrO4 
solution  is  evaporated  to  a  concentration  of  1.5  sp.gr. 

3.  Transformation   of   Na2CrO4   into   Na2Cr207.     If  H2SO4  is 
added  to  the  concentrated,  aqueous  solution  of  the  chromate 
the  following  reaction  takes  place: 

2Na2CrO4  +  H2SO4  =  Na2Cr2O7  +  H2O  +  Na2SO4. 


CHROMIUM 


221 


The  Na2SO4  is  deposited,  to  a  large  extent,  as  a  crys- 
talline powder.  The  liquor  is  drawn  off  from  the  sediment, 
and  the  solution  concentrated,  whereby  practically  all  of  the 
remaining  Na2SO4  separates  out.  If  care  was  taken  to 


FIG.  195. 


p n       n       n a o 0 Q 3 


FIG.  197. 
FIGS.  195-7. — Roasting  Furnace. 

permit  from  i  to  2%  of  neutral  chromate  to  remain  in  the 
solution,  the  evaporation  may  take  place  in  iron  vessels. 
After  the  dehydration  is  almost  complete,  the  melted 
is  poured  into  flat  pans  in  which  it  solidifies. 


222  METALLURGY 

4.  Reduction  of  Na2Cr207  with  S  is  accomplished  by  melt- 
ing the  mixture  in  cast-iron  kettles,  which  are  set  in  masonry 
over  a  grate.  Na2Cr2O7  +  S=Na2SO4  +  Cr2O3.  The  melt  is 
ladled  out,  broken  up  after  it  has  become  cold,  and  leached 
with  water;  the  Na2SO4  passes  into  solution,  while  Cr2O3 
remains  undissolved  and  is  separated  by  decanting  and 
filtering  off  the  solution. 
II.  Reduction  of  the  Cr203. 

By  a  Reducing  Roast.  If  it  is  not  necessary  for  the  Cr  to 
be  obtained  in  a  fused  condition,  it  suffices  to  mix  the  Cr2O3 
with  wood  charcoal,  or  powdered  coke,  and  heat  the  mixture 
in  crucibles.  Even  in  the  regenerative  gas  furnace,  the 
reduced  metal  is  not  fused,  but  is  to  be  found  at  the  bottom 
of  the  crucible  in  the  form  of  a  powder. 

By  a  Reducing  Fusion  with  C.  This  can  be  accomplished  only  in 
electric  furnaces.  It  is  advantageous  to  add  a  little  Al2O3to  the 
mixture  of  Cr2Os  +  3C  and  to  use  as  flux  some  CaF2  or  some 
AlF3.3NaF.  The  A12O3  is  reduced  to  metal  with  the  Cr  at 
the  temperature  required  for  the  melting,  but  it  then  acts  as 
reducing  agent  upon  some  of  the  remaining  Cr2O3. 
APPARATUS:  Heroult  or  Girod  furnaces.  See  Iron,  pp. 

208,  210. 

By  Reducing  Fusion  with  Al.  A  mixture  of  Cr2O3+Al2  is 
kindled  by  an  ignition  powder  of  3BaO2  +  A12;  the  mass  then 
continues  to  fuse  with  great  evolution  of  heat  and  there  is 
formed  A12O3  and  Cr2.  This  is  GOLDSCHMIDT'S  THERMITE 
PROCESS. 
APPARATUS:  crucibles  made  of  MgO  or  else  lined  with 

MgO.  It  is  best  to  embed  the  crucibles  in"  sand. 
Electrolysis  of  Chromium  Solutions.  The  fact  that  chromium 
can  be  deposited  electrolytically  from  aqueous  solutions  of 
chromous  chloride  was  discovered  by  Bunsen  in  1854.  This 
scientist  showed  that  a  relatively  pure  metal  could  be  pro- 
duced by  using  high  current  densities  (at  least  70  amperes 
per  square  foot  of  electrode  surface)  and  concentrated  CrCl2 
solutions.  Both  conditions  are  difficult  to  maintain,  for  the 


CHROMIUM  223 

surface  of  the  electrode  is  constantly  increasing  by  deposition 
of  Cr  and  the  concentration  of  the  solution  is  constantly  diminish- 
ing. For  this  reason  the  author  in  1887  to  1900  used,  instead 
of  the  CrCl2  solution,  a  paste  of  CrF3  crystals  packed  in  a  linen 
bag  and  suspended  in  a  vessel  of  water  through  which  SO? 
was  constantly  passed  during  the  electrolysis  to  effect  depolariza- 
tion. The  cathodes  consisted  of  Pt  foil,  the  anodes  of  C  plates. 
The  current  density  was  gradually  increased  during  the  elec- 
trolysis, to  compensate  for  the  gain  in  electrode  surface.  The 
Cr  grew  upon  the  cathode  foil  in  a  crystalline  condition. 

By  the  more  recent  studies  of  G.  Glaser,  the  limits  of  con- 
centration and  of  current  density  have  been  established  for 
CrCb  solutions. 
ELECTROLYTE:    CrCl2  solutions  containing  13  to  20  oz.  Cr 

per  gallon. 

ANODES:    rods  of  carbon. 
CATHODES:   Pt  foil. 

CURRENT  DENSITY:   85  to  170  amperes  per  square  foot  of 
electrode  surface.     With  densities  of  70  amperes  per  square 
foot,    the    metal   contains   perceptible    amounts   of    CrO, 
and  with  8  amperes  per  square  foot,  only  CrO  is  deposited. 
TEMPERATURE:    50°  C.  (122°  F.),  at  the  most.     At  higher 
temperatures  the  Cr  is  not  deposited  in  a  crystalline  con- 
dition, but  in  the  form  of  a  black  powder. 
Properties  of  Chromium : 
SPECIFIC  GRAVITY:   6  to  7. 
COLOR:  light  gray  with  high  luster. 
MECHANICAL  PROPERTIES  :  hard  and  brittle. 
STRUCTURE:   coarsely  crystalline. 
MELTING-POINT:   1515°  C.  (2760°  F.)  (?). 
BOILING-POINT:   2500°  C.  (4532°  F.). 

ALLOYS  readily  with  Fe,  Mn  and  W;  difficultly  soluble 
in  most  other  metals.  Some  melted  products,  formerly 
regarded  as  alloys,  have  proved  to  be  mixtures  upon  more 
accurate  metallographical  investigation  In  these  the  Cr 
lies  finelv  divided  in  the  other  metal. 


224  METALLURGY 

CHEMICAL  BEHAVIOR:  At  low  temperatures  the  metal  is 
fairly  stable  but  at  high  temperatures  it  combines  energet- 
ically with  most  of  the  metalloids,  particularly  with  C  and 
Si.  It  is  more  readily  dissolved  in  caustic  alkali  solutions 
than  in  acids.  With  the  former,  it  forms  salts  in  which  the 
Cr  is  present  in  the  anion;  with  the  latter,  salts  in  which 
the  Cr  exists  as  a  bivalent  or  trivalent  cation. 


S: 


TUNGSTEN 

Sources. 

Natural  Sources : 
AN  OXIDE  occurs,  called 

Tungstite,  WO3. 
THE  SALTS,  called  tungstates. 

.__.      r  These  tungstates  are  more  widely 
Wolframite,  FeWO4.          ,..,.,  . ,          , 

o  i-    iv      ^  iTr^        1       distributed  than  the  oxide  and 
Scheelite,  CaWO4.  .,    .,    ,  ,    ~.  , 

[      accompany  cassitente  (cf.  Tin). 

Other  Sources: 

SLAGS  obtained  in  smelting  tin  ores  that  contain  tungsten. 

(A)   Ferro-Tungsten. 

In  the  chapter  on  Iron,  it  was  mentioned  that  tungsten  is  used 
in  the  manufacture  of  a  special  kind  of  steel,  and  that  such 
steels  could  be  prepared  by  the  addition  of  an  iron  alloy  rich  in 
W  (80  to  85%  W).  This  alloy,  called  ferro-tungsten,  is  prepared 
by  a 

Reducing  Fusion  of  wolframite,  or  scheelite,  with  powdered 
quartz  and  glass;  these  fluxes  slag  the  bases  other  than  iron 
that  occur  in  the  ores,  particularly  the  alkaline  earths.  In 
working  with  wolframite  it  is  well  to  remember  that  the  ore 
often  contains  considerable  amounts  of  hiibnerite,  MnWO4. 
When  scheelite  is  used,  it  is  obvious  that  Fe  must  be  added  to 
the  charge. 
APPARATUS  AND  OPERATIONS  are  the  same  as  with  ferro 

chrome  (p.  218). 

225 


226  METALLURGY 


(B)  Pure  Tungsten. 

Since  scheelite,  on  account  of  its  low  Mn  content,  is  better 
suited  for  preparing  ferro- tungsten  than  is  wolframite,  which 
almost  always  contains  considerable  Mn,  the  latter  is  used  in 
large  quantities  in  -the  preparation  of  pure  tungsten.  In  working 
up  this  ore,  or  tin  slags  likewise  containing  W  as  tungstate, 
there  are  a  great  many  points  of  resemblance  to  the  process  for 
obtaining  pure  chromium. 
I.  Separation  of  W  from  Fe,  Mn,  Ca,  etc. 

1.  Oxidizing  Roast  with  Alkaline  Fluxes.      Purpose:  transform- 
ation of  W  into  a  soluble  alkali  tungstate,  Na2WO4,  of  Fe 
and  Mn  into    insoluble  oxides,  and   of   Ca   into  insoluble 
carbonate. 

2FeWO4  +  O  +  2Na2CO3  =  2Na2WO4  +  Fe2O3  +  2CO2. 

APPARATUS  AND  OPERATIONS  as  with  Chromium. 

2.  Lixiviation.     Purpose:    separation  of    the  soluble  tungstate 
from  the  substances  that  are  insoluble  in  water.    As  solvent, 
the  weak  wash  waters  of  the  previously  roasted  product  are 
used.     The  roasted  product,  as  it  comes  from  the  furnace, 
is  leached    hot    in    this   liquor.     By  thus  quenching  it,  the 
roasted  product  is  broken  up  where  it  has  been  sintered  and 
put  into  a  condition  more  favorable  for  leaching.     When 
the  solution  has  a  concentration  of  10  to  12%  tungstate, 
it  is  drawn  off  and  concentrated  by  evaporation.     Hereby 
certain  contaminating  salts,  such  as  Na2SO4,  which  dissolve 
in  water  with  the  tungstate,  separate  out.     When  the  solution 
is  sufficiently  concentrated  it  is  cooled  and  brought  to  crystal- 
lization. 

3.  Precipitation  of  the  So-called  Tungstic   Acid,  W03.     If  the 
concentrated  solution  of  tungstate  is  heated  by  introducing 
steam,   and  then  hydrochloric  acid  is  allowed  to  run  in, 
or  if  the  powdered  crystals  of  tungstate  are  shaken  into  the 


TUNGSTEN  227 

hot  solution  of  hydrochloric  acid,  the  WOs  is  deposited  as  a 
heavy,  yellow  powder  which  can  be  purified  by  decantation, 
nitration,  and  drying. 

Na2  WO4  +  2HC1  =  2NaCl  +  H2O  +  WO3 

The  precipitation  and  first  washing  take  place  in  stone  vats, 
the  last  washing  in  filter  bags. 
II.  Reduction  of  W03. 

Reduction  without  Fusion.  Mixtures  of  about  265  Ibs.  WOs,  26 
to  33  Ibs.  wood  charcoal  or  coke,  and  15  to  9  Ibs.  of  powdered 
colophonium,  or  pitch,  are  placed  in  crucibles  and  heated 
as  hot  as  possible  in  gas  furnaces  with  blast  or  regenerative 
firing.  In  this  case  the  reduced  metal  is  not  melted. 
APPARATUS  AND  OPERATIONS  as  in  the  reduction  of  Chromium 

(p.  222), 

Reducing  Fusion.     If  the  same  charge  as  used  for  the  above 
reduction  is  placed  in  an  electric  furnace,  some  difficulty  will 
be  experienced  in  fusing  it.     The  melting  point  of  W  lies  very 
high   (2800°   to   2850°   C.)   and  although  a  temperature  of 
3500°  C.  is  reached  in  the  electric  arc,  this  region   of  high 
temperature  is  of  very  limited  area.     Hence  the  electrodes 
must  be  brought  as  close  to  the  metal  as  possible,  or  else 
the  latter  must  be  made  one  pole  in  the  circuit  of  the  arc, 
and  by  so  doing  it  is  very  hard  to  avoid  the  taking  up  of  C. 
Carbon  is  dissolved  with  the  formation  of  the  carbides  W2C 
and  WC.     Metal  containing  carbide  can  be  fused  readily 
in  the  electric  arc. 
Properties  of  Tungsten : 
SPECIFIC  GRAVITY:  19. 
COLOR:   gray,  crystalline  as  powder;   in  the  fused  condition 

it  is  a  nearly  white,  lustrous  metal. 
MECHANICAL  PROPERTIES  :  Very  hard. 
MELTING  POINT:  2800°  to  2850°  C.  (5070°  to  5160°  F.). 
BOILING  POINT:  3700°  C.  (7000°  F.). 
ALLOYS  with  other  metals  to  about  the  same  degree  as  Cr  does. 


228  METALLURGY 

CHEMICAL  BEHAVIOR:  It  is  oxidized  by  O  only  at  high 
temperatures;  likewise  the  halogens  act  vigorously  only 
when  the  metal  is  hot.  It  is  insoluble  in  most  acids. 
Strongly  oxidizing  acids  produce  WOs;  this  oxide  is  an 
acid  anhydride  and  combines  readily  with  basic  oxides 
to  form  tungstates. 


CADMIUM 

Sources 

Natural  Sources : 

THE  SULPHIDE,  CdS,  called  greenockite,  is  of  rare  occurence 
and  is  usually  accompanied  by  ZnS  in  the  blendes  of 
Silesia  and  of  North  America. 
THE  CARBONATE,  CdCO3,  is  found,  as  a  rule,  only  with  the 

smithsonites  of  Silesia  and  of  North  America. 
Other  Sources:     The  zinc  dust  formed  in  working  up  zinc 
ores  containing  cadmium. 

(A)  Extraction 

Since  there  are  not  enough  cadmium  ores  available  to  support 
an  independent  industry,  the  production  of  cadmium  forms 
merely  a  side-issue  in  zinc  smelters  that  work  zinc  ores  con- 
taining cadmium.  As  will  be  seen  under  the  section  on  Zinc 
(p.  234),  when  such  ores  are  roasted  CdO  is  formed  together  with 
ZnO  and  it  is  the  former  oxide  that  is  first  reduced  to  metal. 
The  Cd  is  found,  therefore,  partly  as  metal,  and  partly  as  carbonate 
and  oxide,  together  with  large  amounts  of  Zn  and  ZnO  (70  to 
80%)  in  the  first  zinc  dust  that  is  obtained.  The  preparation  of 
Cd  from  this  raw  material  consists  in  a  repeated 

Reduction  with  Fractional  Distillation,  first  in  clay  and  finally 
in  iron  retorts.  The  method  of  working  is  the  same  as  for 
the  reduction  of  ZnO,  except  that  the  temperature  is  kept 
lower.  For  details,  see  Zinc. 

229 


230  METALLURGY 


(B)  Refining 

The  crude  cadmium,  containing  more  or  less  zinc,  can  be  puri- 
fied further  by  repeating  the  above-mentioned 

Reduction    with    fractional    distillation    until    finally   the  Zn 
content   is  brought  very  low.     It  can  also  be   freed  from 
Zn  by 
Electrolysis. 

ANODES  :  crude  Cd  containing  Zn. 
CATHODES  :   pure  Cd. 
ELECTROLYTE:   CdCl2  or  CdSO4. 
CURRENT  DENSITY:    6  to  15  amperes  per  square  foot. 
E.M.F.:  2.8  to  3.5  volts. 
Properties  of  Cadmium  : 

SPECIFIC  GRAVITY:  8.6  to  8.7. 
COLOR:  white,  strongly  lustrous. 
MECHANICAL  PROPERTIES:  soft,  ductile. 
STRUCTURE:    It  shows  a  tendency  to  form  dendrites  on  the 
surface,    otherwise    granular.      The    crystal    grains    grow 
considerably  upon  long-continued  heating. 
ALLOYS  with    most    other    metals.      It    forms    some    alloys 
with  remarkably  low  melting-points;    e.g.,  4  pts.  Bi,  i  pt. 
Sn,  2  pts.  Pb,  i  pt.  Cd  =  Wood's  Metal.     15  pts.  Bi,  4  pts. 
Sn,  8  pts.  Pb,  3  pts.  Cd  =  Lipowitz  Metal.     The  former 
alloy  melts  at  71°  C.   (160°  F.)  and  the  latter  at  60°  C. 


CHEMICAL  BEHAVIOR  :  At  low  temperatures  it  is  stable  in  the 
air  but  unites  readily  with  the  halogens.  At  high  tempera- 
tures it  burns  readily  in  O  or  S.  It  is  readily  soluble  in 
HC1,  H2SO4,  and  HNO3,  also  in  alkali  hydroxides.  E.M.F. 
toward  H=  +0.420  volt. 


ZINC 

Sources 

Natural  Sources : 

ZINC  BLENDE,  or  sphalerite,  ZnS.  Gangue  and  accompanying 
minerals  are 

1.  Galena  in  Germany  in  the  ore  deposits  at  Ems  and 
the  Upper  Harz,  also  in  Bohemia  and  Hungary,  and 
in  Australia. 

2.  Dolomite,    limonite,    and    also    smithsonite,    often    in 
clay,  in  Rhineland,  Westphalia,  Belgium,  Upper  Silesia, 
Northern  Spain,  Algiers,  and  North  America. 

3.  In  gneiss  with  pyrites  in  Sweden. 

SMITHSONITE*  or  Zinc  Spar,  ZnCO3.  Gangue:  often  with 
sphalerite  and  galena,  always  with  zinc  silicates  in  lime- 
stone, dolomite  and  also  limonite  in  Rhineland,  Belgium, 
North  Spain,  England,  North  America. 

CALAMINE,  H2Zn2SiO5. 

WILLEMITE,  Zn2SiO4.  The  gangue  and  accompanying  miner- 
als are  the  same  as  with  smithsonite. 

RED  ZINC  ORE  or  zincite  ZnO.  Gangue;  the  same  as  with 
franklinite. 

FRANKLINITE:  (Zn,  Fe,  Mn)  0  (Fe2,  A12,  Mn2)  O3.     It  is 
found  in  New  Jersey  near  the  village  of  Franklin  (whence 
the  name)  associated  with  zincite,  willemite,  rhodonite,  and 
tephroite,  etc. 
Other  Sources : 

ZINC  DUST:  the  first  condensation  product  in  the  reduction 
of  zinc  with  sometimes  as  much  as  90%  Zn. 

*  Sometimes  called  calamine,  but  incorrectly. 

231 


232  METALLURGY 

FLUE  DUST  from  smelting  furnaces. 

FURNACE  CALAMINE,  the  deposit  in  shaft  furnaces  in  which 

ores  containing  Zn  are  smelted. 
DROSS,  waste  metal  obtained  in  the  refining  of  zinciferous 

metals,   and   waste   from   foundries,   plating   works,    etc. 
ZINC  SKIMMINGS,  alloys  of  Ag,  Pb,   Cu  and  Zn  obtained 

in  the  removal  of  Zn  and  A    from  lead  bullion. 


(A)  Extraction  of  Crude  Zinc 

Roast  Reduction  Work  : 

i.  The  Roasting  effects  the  transformation  of  ZnS  and  other 
sulphides,  as  well  as  of  ZnCO3  and  other  carbonates,  into 
oxide.  Any  water  held  mechanically  or  chemically  is  also 
driven  out. 

The  principal  reactions  that  take  place  are 

ZnCO3  = 


As  compared  with  the  other  sulphides,  sphalerite  is  diffi- 
cult to  roast;  basic  sulphates  as  well  as  oxide  are  formed 
and  the  former  are  hard  to  decompose.      Both    the  oxide 
and  the  basic  sulphates  form   a  coating  over  the  sulphide 
and  prevent  the  access  of  air.    A  complete  roasting,  which 
is  necessary  on  account  of  the  injurious  action  of  sulphides 
upon  the  retorts,  can  be  accomplished  only  with  the  applica- 
tion of  external  heat. 
ROASTING  APPARATUS: 
HEAPS  AND  OPEN  KILNS  for  drying,  and  rarely  for  calcining, 

material  containing  carbonates. 
SHAFT  FURNACES  for  burning  carbonate  material. 
REVERBERATORY    FURNACES    with    fixed    hearths,    worked 

by  hand  or  in  some  cases  provided  with  mechanical  stirrers. 

Recently  for  roasting  blende  that  contains  pyrite,  rever- 

beratory  furnaces  with  revolving  hearths  have  been  some- 

times used. 


ZINC 


233 


MUFFLE  FURNACES,  first  introduced  by  Liebig  and  Eichhorn 
by  arranging  heating  flues  in  the  fine  pyrites  burner  of  the 
Maletra-Schaffner  System,  and  then,  with  success  in  zinc 
smelters,  by  Hasenclever.  In  furnaces  of  this  type,  four 
men  (2  for  each  12  hours)  can  roast  from  4  to  4.5  tons  of 


X&&&-  ^iwi? 


FIG.  199. 
Hasenclever  Furnace. 

ore  in  a  day.     From  300  to  450  Ibs.  of  coal  are  required 

per  ton  of  ore. 

2.  Reduction  of  the  Roasted  Product.  To  carry  out  the 
work  to  the  best  advantage,  the  properties  of  zinc  oxide 
and  of  zinc  necessitate  the  removal  of  the  metal  from  the 
reduction  apparatus  in  the  form  of  vapor  and  the  condensa- 


234  METALLURGY 

tion  of  the  vapor  in  receivers;  in  other  words  a  distillation 
is  combined  with  a  reduction.  Although  the  reduction 
of  zinc  oxide  by  carbon  will  take  place  at  a  distinct  red  heat, 
still  the  reaction  under  these  conditions  is  so  slow  and  incom- 
plete that  to  obtain  satisfactory  results  the  reduction  tem- 
perature of  the  zinc  oxide  must  be  kept  between  1000°  and 
1300°  C.;  the  melting-point  of  Zn  lies  at  415°  C.  and  the 
boiling-point  at  930°  to  950°  C. 

FLUX:  An  abundance  of  carbon  to  prevent  the  formation 
of  CO?  during  the  reduction;  this  gas  will  oxidize  Zn  to 
ZnO  at  red-heat  in  the  receivers  and  results  in  the  forma- 
tion of  zinc  dust. 

Zinc  silicates  do  not  require  the  addition  of  any  flux  to 
combine  with  the  silica,  for  the  ZnO  contained  in  the 
silicate  is  reduced  and  the  SiO2  remains  behind  as  such 
or  in  the  form  of  acid  silicates,  and  tends  to  make  the  residue 
fusible;  the  more  difficultly-fusible  the  residue  the  better 
for  the  apparatus  in  which  the  work  is  conducted. 
BEHAVIOR  OF  THE  CONSTITUENTS  OTHER  THAN  ZnO  that 
are  present  in  the  charge:  Zinc  sulphide  is,  to  be  sure, 
not  reducible  by  carbon,  but  usually  iron  compounds  are 
present  with  which  it  reacts  to  form  FeS  and  Zn;  FeS 
has  a  very  injurious  action  upon  the  walls  of  the 
apparatus. 

Cadmium  compounds  behave  similarly  to  the  corre- 
sponding zinc  compounds  except  that  CdO  is  more  readily 
reducible  and  melts  at  a  lower  temperature;  it  passes  over, 
therefore,  with  the  first  of  the  zinc  dust. 

Of  the  iron  and  manganese  compounds,  Fe2O3  and 
Mn2Oy  are  reduced  to  FeO,  Fe,  and  MnO.  FeO  and 
MnO  unite  with  silicates  to  form  readily-fusible  slags 
and  these  are  dangerous  for  the  retorts.  FeS  is  unacted 
upon,  but  melts  and  injures  the  walls. 

Lead  oxide  in  the  free  state  is  easily  reduced  and  is 
volatilized  to  some  extent  with  the  Zn;  combined  with 
silica,  it  forms  fusible  silicates  which  injure  the  walls. 


ZINC  235 

Free   silica  is  useful,   since   the    melting-point    of    the 
residue  increases  in  proportion  to  the  SiC>2  content. 
Under  the  different  modes  of  working  that  prevail  in  dif- 
ferent  countries   and   in   connection   with   different   mines, 
the  usual  apparatus  and  processes  followed  may  be  classified 
into  three  groups:   the  Silesian,  the  Belgian,  and  the  Rhenish 
methods. 

In  all  cases,    muffles,  or  tube- shaped  retorts,  closed  at  one 
end,  are    used  as  reduction  vessels.     The  retorts  are  either 
oval,  semi-oval,  or  cylindrical  in  form.     The  size  and  shape 
of  the  retort  are  regulated  by  the  following  circumstances. 
ZINC  CONTENT:    the  greater  the  zinc  content,    the   smaller. 

the  retort. 
REDUCTIBILITY :  the  harder  it  is  to  reduce  the  ore,  the  smaller 

the  retort. 

FUEL:  the  influence  of  the  fuel,  now  that  gas  furnaces  are 
used  almost  universally,  steps  into  the  background. 
Formerly,  furnaces  with  a  great  many  small  retorts  required 
a  fuel  that  would  give  a  long  flame,  and  furnaces  with 
large  retorts  could  get  along  with  a  short-flame  fuel. 
These  factors  have  led  to  the  development  of  the  following 
systems  of  retorts  with  furnaces  suitable  for  them: 
In  Silesia,  where  the  ores  are  poor  and  where  to  some  extent 
coal  giving  short  flame  is  available,  the  work  is  carried 
out  in  large  muffles.  The  Silesian  muffles  have  the  fol- 
lowing inside  dimensions:  height  26  in.,  breadth  6  to  8 
in.;  length  27  to  85  in.;  thickness  of  walls,  f  in.  in  front 
and  2\  in.  in  back. 

THE  CONDENSERS  for  Silesian  muffles  are  metal  tubes, 
metal  boxes,  and  also  clay  tubes,  conical  or  bellied  in  form. 
CAPACITY:    up  to  220  Ibs.  ore,  together  with  about  40% 
as  much  of  reduction  carbon. 

ARRANGEMENT  IN  THE  HEATING  CHAMBERS:  in  a  row, 
as  many  as  36  pieces  side  by  side;  two  heating  chambers 
lie  with  the  vertical  back  walls  opposite  one  another  so  that 
one  furnace  may  hold  as  many  as  72  muffles. 


236 


METALLURGY 


FIRING:   usually  regenerative. 

CONSUMPTION  OF  FUEL:  one  Silesian  furnace  with  60 
muffles,  each  of  220  Ibs.  capacity,  will  work  up  6  tons  of  ore 
in  24  hours,  using  thereby  6.1  tons  of  heating  carbon  and  2.4 
tons  of  reducing  carbon. 

YIELD:    i  ton  of  zinc;  hence  6.1  tons   of  heating  carbon 
and  2.4  tons  of  reducing  carbon  are  used  per  ton  of  zinc. 
IN  BELGIUM  where  the  ores  are  rich  and  long-flame  coal 

is    available,    small   retorts   are  used  which    are    either 


FIG.  200. — Silesian. 


FIG.  201. — Rhenish. 


FIG.  202. — Belgian. 
Types  of  Muffles.     Scale  i :  50. 

cylindrical  tubes  6  to  10  in.  in  diameter,  or  oval  tubes 
6J  to  7  in.  wide  and  7!  to  n  in.  high;  these  retorts  are  40 
to  57  in.  long  and  the  walls  are  f  in.  thick  in  front  and 
i  J  in.  thick  in  the  rear. 

CONDENSERS  FOR  THE  BELGIAN  RETORTS:  conical  clay 
tubes  1 6  in.  long,  and  with  a  diameter  of  3  in.  in  front  and 
6  in.  in  back.  The  wide  end  is  shoved  into  the  retort 
and  the  narrow  end  is  fitted  into  a  sheet-iron  condenser, 
or  nozzle,  which  serves  to  catch  the  zinc  dust. 


ZINC 


237 


238 


METALLURGY 


CAPACITY  OF  THE  BELGIAN  RETORTS:    about  44   Ibs. 
ore  with  40  to  60%  as  much  reducing  carbon. 


ARRANGEMENT  OF  THE  BELGIAN  RETORTS  IN  THE 
HEATING  CHAMBERS:  the  retorts  are  placed  one  over 
another  in  six  or  eight  rows;  every  two  or  three  vertical 


ZINC 


239 


•series  open  into  a  common  niche  in  which  the  condensers  lie. 
The  back  end  of  the  tube  rests  against  a  supporting  wall, 
and  the  front  end  in  the  wall  of  the  niche;  the  rest  lies 
free.  In  one  furnace-block,  with,  as  a  rule,  the  back 
walls  of  two  adjacent  heating  chambers  lying  against  one 
another,  there  are  from  56  to  400  of  these  tubes  or  retorts. 
FIRING:  Siemens'  regenerative  or  recuperative  firing 
of  different  systems;  direct  coal  firing  is  seldom  used  in 
modern  plants;  in  the  United  States  natural  gas  is  also 
^employed  as  fuel. 


FIG.  205. — Part  of  Silesian  Furnace. 

CONSUMPTION  OF  FUEL:  furnaces  with  56  or  70  retorts 
require  respectively  for  i.o  and  1.35  to  1.44  tons  of  ore, 
1.8  and  2.5  to  2.55  tons  of  fuel. 

YIELD:  One  ton  of  zinc  produced  required  4.5  to  5.5 
tons  of  fuel,  and  1.3  to  1.9  tons  of  reducing  carbon. 
IN  ZINC  SMELTERS  OF  THE  RHENISH  PROVINCES  the  attempt 
has  been  made  to  combine  the  advantages  of  large  muffles 
with  the  possibility  of  arranging  them  as  is  customary 
with  the  smaller  retorts.  Semi-oval  retorts  are  used,  varying 
in  size  between  the  Silesian  muffles  and  Belgian  retorts, 
and  so  chosen  that,  with  the  full  burden,  they  can  be  placed 
free  in  the  heating  chamber  according  to  the  Belgian 
method. 


240 


METALLURGY 


ZINC 


241 


The  dimensions  of  the  retorts  in  common  use  in  the 
Rhenish  provinces  are:  breadth  up  to  6\  in.,  height  up  to 
i if  in.,  length  up  to  55  in. 

CONDENSERS  FOR  THE  RHENISH  RETORTS:  bulging 
clay  tubes  open  at  both  ends,  of  which  the  wider  end  is 


FIG.  207. — Belgian  Furnace  with  Siemens,  Regenerative  Chambers. 

inserted  into  the  end  of  the  retort  and,  at  the  beginning 
of  the  work,  the  front  ends  are  stuck  into  the  sheet  iron 
prolongs,  so  as  to  catch  the  zinc  dust. 

CAPACITY  OF  THE  RHENISH  RETORTS:  55  to  75  Ibs.  of 
ore  with  40  to  50%  as  much  reducing  carbon. 

ARRANGEMENT  OF  THE  RETORTS  IN  THE  HEATING 
CHAMBERS:  the  retorts  rest  above  one  another  in  three 


242  METALLURGY 

horizontal  layers;  every  two  vertical  columns  open  into 
a  front  niche  in  which  are  the  receivers.  The  retorts 
lie  as  in  the  Belgian  type.  The  furnace  block  may  con- 
tain 40  to  80  retorts  in  three  series,  i.e.,  120  to  240  retorts 
in  all. 


FIG.  208. — Rhenish  Furnace  with  Recuperative  System. 

FIRING:  Siemens'  regenerative  firing  is  less  used  than 
the  so-called  recuperative  firing. 

CONSUMPTION  OF  FUEL:  a  large  furnace  of  this  type 
with  240  retorts  and  recuperative  firing  requires  8.5  to 
9.2  tons  of  fuel  and  3.2  tons  of  reducing  carbon  in  24  hours 
for  operating  with  8  tons  of  ore  with  52  to  54%  of  zinc. 


ZINC 


243 


YIELD:    2.5  tons  of  fuel  and  i  ton  of  reducing  carbon 
are  required  to  produce  one  ton  of  zinc. 


FIG.  209. — Retort  Press  (C.  Mehler,  Aachen).     Scale  i :  40. 

THE  MAKING  OF  THE  RETORTS  is  carried  out  as  a  rule  at 
the  smelter.  The  work  consists  of  the  following  operations : 
BURNING  of  a  part  of  the  clay  that  is  to  be  used  in  making 

the  retorts. 


244  METALLURGY 

GRINDING. 

PULVERIZING  the  burnt  clay. 

MIXING  the  powdered  clay  with  50  to  100%  as  much  fresh 
clay. 

STAMPING  AND  PRESSING  the  mixture  into  retorts  after 
it  has  stood  for  some  time. 

DRYING  AND  BURNING  THE  RETORTS:  Enough  retorts 
must  be  kept  in  the  burning  furnace  to  replace  those 
that  are  injured  during  the  reduction  of  the  zinc  so 
as  to  avoid  the  dangeir  of  injury  that  would  be  likely 
to  take  place  in  the  cooling  and  reheating.  The  hot 
retorts,  therefore,  are  introduced  from  this  furnace 
directly  into  the  retort-furnace  as  they  are  required. 

LIFE  OF  MUFFLES.  The  retorts  last,  according  to  the 
amount  of  work  done  in  the  smelter,  from  14  to  30  days; 
this  corresponds  to  a  daily  replacement  of  3  to  7%. 

It  is  worth  mentioning  that  many  experiments  have  been 
carried  out  in  the  attempt  to  utilize  electricity  for  heating  in  the 
roast-reduction  work.  Thus  the  Cowles  Brothers  in  1885 
proposed  the  use  of  retorts  which,  by  having  carbon  contacts 
at  both  ends,  were  connected  with  an  electric  circuit  using  the 
charge  as  the  heat-producing  resistance.  The  constant  changes 
taking  place  in  the  body  of  the  charge  by  the  reduction  process 
and  the  fact  that  as  the  temperature  rises  the  retort  and 
the  walls  take  part  more  and  more  in  the  conduction  of  the 
current,  have  given  rise  to  difficulties  which  have  not  yet  been 
overcome. 

The  Precipitation  Process,  consisting  of  the  smelting  of  sulphide 
ores  with  iron  or  fluxes  containing  iron,  has  been  tested  in 
different  directions. 

In  the  blast-furnace,  the  possibility  of  such  a  process  is 
altogether  out  of  the  question,  because  the  reaction  gases  can 
never  be  kept  absolutely  free  from  SO2  and  CO2.  Thus  the  Zn 
would  never  be  obtained  as  a  fused  metal. 

Experiments  in  which  the  ZnS  was  exposed  to  the  action  of 


ZINC  245 

molten  Fe  in  a  rotary,  closed  converter,  also  proved  fruitless. 
The  requisite  mixing  of  the  ZnS  and  Fe,  which  is  necessary 
for  the  rapid  fulfilment  of  the  desired  reaction,  could  not  be 
accomplished  by  revolving  the  drum. 

Apparently  the  problem  has  been  solved,  however,  by  heating 
a  mixture  of  finely-divided  ZnS  with  equally  fine  Fe  in  an  elec- 
tric furnace  in  which  a  bath  of  FeS  is  formed  b  the  reaction 


and  this  kept  as  a  heating   resistance.     The  FeS  holds  Fe  as 
well  as  ZnS  in  solution,  so  that  it  serves  to  maintain  the  desired 
state  of  subdivision,  and   the    reaction    between  the  Fe  and 
ZnS  continues  to  take  place  smoothly. 

The  Direct  Electrolytic  Process  for  working  up  crude  zinc  from 
ores  by  making  the  latter  the  anode  in  an  aqueous  solution 
of  zinc  salt,  has  never  met  with  success. 

(B)    Zinc  Refining 

By  Liquating  in  reverberatory  furnaces,  in  which  the  bottom  of  the 
hearth  falls  away  from  the  fire-bridge  toward  the  opposite 
wall,  it  is  customary  to  purify  the  crude  zinc,  obtained  from 
the  Roast-Reduction  Process,  which  usually  contains  Pb.  From 
the  fused  zinc  a  specifically  heavier  alloy  of  Pb  or  Fe  with  Zn 
separates  out  at  the  bottom  and  is  removed  from  time  to  time 
from  the  deepest  part  of  the  hearth.  On  top  of  the  metal  bath 
float  the  more  infusible  impurities  which  were  mechanically 
enclosed  by  the  crude  zinc;  these  are  skimmed  off.  The  zinc 
itself  is  usually  dipped  out  with  iron  ladles,  poured  into  plates, 
and  rolled  at  a  suitable  temperature. 

Distillation  is   employed   with   alloys  rich    in   lead,  containing 
precious  metals  (skimmings  rich  in  Zn  from  the  Zn-desilveriza- 
tion). 
APPARATUS: 

Crucibles   in  air-furnaces;    this   is  the  oldest  process  for 
small   works. 


246  METALLURGY 

Bottle-shaped    distilling   vessels    in    tilting     air-  furnaces, 

(Figs.  210  to  212.) 

Bottle-shaped  or    tube-shaped    distilling  apparatus  made 
of  fire-clay  with  graphite  lining,  heated  in  reverberatory 
furnaces.    (Figs.  213  and  214.) 
Electrolytic  Separation  of  Zn  alloys  has  not  met  with  success 

in  practice  (e.g.,  for  working  up  skimmings  rich  in  Zn). 
Lixiviation  and  Electrolytic  Deposition  has  been  attempted  at 
some  places  without  meeting  with  commercial  success,  but  a 
practical  process  has  been  worked  out  at  the  works  of  Brunner, 
Mond  &  Co.,  at  Norwich,  England,  using  the  patents  of  Hopfner. 
It  consists  of  the  following  operations  : 
Chloridizing    Roast  of  Zn  ores,  or  pyritic  residues  containing 

Zn,  in  reverberatory  furnaces. 
Lixiviation  with  water. 
Purification  of  the  liquor  by 

CRYSTALLIZING  out  of  the  sulphate  or  decomposing  it  with 

CaCl2. 
PRECIPITATION  of  Fe  and  Mn  by  means  of  Ca(OCl)2. 


PRECIPITATION    of    the    more    electro-positive    metals    with 

zinc  dust. 

Electrolysis  of  the  Liquor. 
ANODES:  carbon  plates. 
CATHODES  :  rotating  Zn  plates. 
ELECTROLYTE:    ZnCl2   solution   with   0.08   to   0.12%    free 

HC1. 

CURRENT  DENSITY:    10  amperes  per  square  foot. 
E.M.F.:  3.3  to  3.8  volts. 
THE   APPARATUS   is   very  complicated,  because   the   anode 

plates  must  be  placed  in  special  cells,  and  each  be  provided 

with  piping  for  carrying  away  the  chlorine  set  free  during 

electrolysis. 

Although  the  practical  results  have  never  been  encouraging, 
the  sulphatizing  roast  of  Zn  ores  would  seem  more  advantageous 


ZINC 


247 


0 

0               C 

5               0 

0 

CD 

n 

H 

3Hr 



C2 

i  —  i 

o 

0              C 

)              0 

o 

248  METALLURGY 

as  the  sulphate  liquors  can  be  electrolyzed  without  the  use 
of  diaphragms.    Moreover,  useful  by-products  can  be  produced 
at  the  anodes  and  in  this  way  the  E.M.F.  may  be  utilized  to 
the  best  advantage. 
Properties  of  Zinc : 

SPECIFIC  GRAVITY:   6.9  to  7.2. 
COLOR:  bluish- white  with  metallic  luster. 
MECHANICAL    PROPERTIES:     very    ductile    at    temperatures 
between  100°  and  150°  C.,  can  be  rolled  into  sheets,  ham- 


FIG.  215. — Structure  of  Bar  Zinc. 

mered  and  pressed;    at  200°  C.  it  becomes  so  brittle  that 

it  can  be  pulverized. 
STRUCTURE:  there  is  a  formation  of  dend rites  on  the  surface 

of  the  metal,  but  the  interior  is  coarse  grained. 
MELTING-POINT:  419°  C.  (786°  F.) 
BOILING-POINT:  940°  C.  (1724°  F.) 
ELECTRICAL  CONDUCTIVITY:  0.27  that  of  Ag. 
ALLOYS  with  most  of  the  metals.     The  alloys  with  Au,  Ag 

and    Pb    (zinc    desilverization)     have     been    mentioned. 


ZINC  249 

Furthermore  there  are  many  alloys  in  practical  use,  such 
as  those  with  Cu  (brass),  with  Cu  and  Sn  (bronze),  with 
Cu  and  Ni  (German  silver),  and  with  Sn  and  Sb  (bearing 
metal). 

CHEMICAL  BEHAVIOR:  Zinc  is  oxidized  superficially  when 
exposed  to  the  atmosphere,  but  the  layer  of  oxide  is  thin 
and  protects  the  metal  within  so  that  zinc  objects  made 
of  thin  foil  will  keep  for  a  very  long  time.  It  is  attacked  by 
the  halogens  at  ordinary  temperature  and  by  the  other 
metalloids  at  higher  temperatures.  The  metal  will  reduce 
H2O  at  a  moderate  red  heat.  On  account  of  its  high 
electrolytic  solution  tension  (E.'M.F.  toward  11=4-0.770 
volts),  it  is  easily  soluble  in  HC1,  H2SO4,  HNO3,  and  in 
caustic  alkali  solutions.  With  the  acids  it  forms  salts  in 
which  the  metal  is  present  as  a  bivalent  cation;  with  the 
alkalies  it  forms  zincates  in  which  the  Zn  is  in  the  anion. 


MANGANESE 

Sources 

Natural  Sources: 

As  OXIDES  AND  HYDRATED  OXIDES  in  the  following  ores: 
Braunite,  the  normal  manganic  oxide,  Mn2Os. 
Hausmannite,    mangano-manganic    oxide,    Mn3O4    (this 
can  be  regarded  as  Mn2MnO4>  manganous  manganite). 
Pyrolusite,  black  oxide  of  manganese,  manganese  peroxide, 
or  dioxide,  MnO2,  the  most  widely  distributed  ore  of 
manganese. 

Manganite,  Mn2O2(OH)2. 
SALTS: 

Rhodocrosite,  or  manganese  spar,  MnCO3.     This  is  an 
important  ore  used  in  the  preparation  of  Mn-Fe  alloys. 
Rhodonite,  MnSiOa- 
Other  Sources: 

SLAGS:  particularly  from  pig  iron  mixers  and  the  prepara- 
tion of  alloys  rich  in  Mn. 
METAL  WASTE,  ferro -manganese. 

(A)   Manganese  Alloys 

The  large  amounts  of  Mn  used  in  the  metallurgy  of  iron,  together 
with  the  fact  that  it  is  difficult  to  prepare  from  pure  ores  by  the 
relatively  inexpensive  smelting  processes  a  pure  metal  that  is 
stable  in  the  air,  make  the  most  important  work  of  manganese 
smelting  the  preparation  of 

250 


MANGANESE  251 

Iron-manganese  Alloys: 

Spiegeleisen,  a  name  given  to  alloys  with  between  5  to  20%  Mn. 

Ferromanganese,  which  includes  the  richer  alloys  with  30  to 
80%  Mn.  The  preparation  of  either  of  these  alloys  takes 
place  by  a  reducing  smelt,  such  as  was  described  under 
Iron,  of  iron  ores  containing  manganese,  or  of  manganese 
and  iron  ores  together  in  a  blast  furnace.  The  working 
conditions  are  changed  only  in  so  far  as  is  necessitated  by 
the  fact  that  as  the  Mn  content  of  the  required  alloy  increases, 
the  temperature  of  the  furnace  and  the  content  of  bases 
in  the  slag  (particularly  of  MnO)  must  be  raised.  The  latter 
contains  from  8  to  30%  MnO.  High  slag  losses  of  the 
Mn  are  unavoidable,  because  if  a  poorer  slag  were  formed, 
the  primarily  reduced  Mn  would  react,  as  it  always  does, 
with  the  oxides  of  iron  in  the  ore.  Furthermore,  large  losses 
of  Mn  occur  through  volatilization. 


(B)    Manganese  Compounds 

Among  the  manganese  compounds  which  come  into  consider- 
ation for  metallurgical  purposes,  or  which  can  be  prepared  in 
smelters,  the  only  ones  to  be  mentioned  are, 

Manganese  Silicides,  or  alloys  of  manganese  silicide.  One  of 
these  silicides  has  the  composition  Mn2Si  and  contains 
approximately  20%  Mn.  There  is  only  one  commercial 
method  for  preparing  silicides,  and  this  consists  of  a  reducing 
smelt  of  manganese  oxides  with  sand  in  an  electric  furnace. 
Since,  however,  most  Mn  ores  contain  enough  Fe  to  make 
the  melted  silicide  unsuitable  for  some  purposes,  such  ores 
are  preferably  smelted  first  for  a  ferromanganese  and  a  slag 
free  from  Fe  which,  with  the  addition*  of  sand,  will  yield 
a  silicide  practically  free  from  Fe  with  from  20  to  22% 
Mn. 


252  METALLURGY 


(C)   Pure  Manganese 

Pure  manganese  cannot  be  obtained  by  smelting  the  purest 
manganese  oxides  with  C.  Even  when  the  amount  of  C  in  the 
charge  is  most  carefully  restricted,  the  metal  obtained  will  contain 
Mn3C  and  as  this  compound  is  decomposed  by  water  at  ordinary 
temperatures,  it  is  not  stable  in  moist  air.  A  fairly  stable  metal, 
low  in  carbon,  but  one  which  is  not  perfectly  pure,  can  be  obtained 
by  the 

Thermite     Process    of    Goldschmidt.     Mn2Os   is   mixed   with 
A12  and  the  reaction  started  by  means  of  ignition  powder 
(3BaO2  +  Al2).     The  heat  of  reaction  is  so   great  that  the 
reduced  Mn  is  entirely  liquefied  and  the  slag  of  A12O3  sepa- 
rates out  of  itself. 
Properties  of  Manganese : 
SPECIFIC  GRAVITY:  7  to  8. 
COLOR:    white,  lustrous. 

MECHANICAL  PROPERTIES:  very  hard  and  brittle. 
FRACTURE  :  smooth,  almost  vitreous. 
MELTING-POINT:    1233°  €.(22 51°  F.) 

VOLATILIZES  considerably  even  at  the  melting  temperature. 
BOILING-POINT:    2200°  €.(4000°  F.) 
ALLOYS  with  Fe,  and  very  readily  with  Cu  and  the  precious 

metals. 

CHEMICAL  BEHAVIOR:  Mn  is  one  of  the  metals  that  enters 
most  readily  into  chemical  reactions.  It  has  a  great  affinity 
toward  almost  all  of  the  metalloids  and  possesses  a  fairly 
high  electrolytic  solution  tension,  both  in  aqueous  solutions 
of  acids  (  +  1.075  volts  toward  H),  as  well  as  in  the  melts 
that  come  into  consideration  in  metallurgical  work  (a  good 
desulphurizing  agent  for  pig  iron). 

Mn  forms  a  series  of  six  oxides  with  corresponding  salts. 
Those  oxides  with  the  least  O  possess  basic  properties, 
whereas  those  containing  more  O  are  acid  anhydrides. 


ALUMINIUM 

Sources 

Natural  Sources : 

OXIDES  AND  HYDRATED  OXIDES. 

CORUNDUM,   pure   A12O3.     The   ruby   and   sapphire   are 

varieties  of  corundum. 
EMERY,  impure  A12O3. 

BAUXITE,  hydrargillite,  A1(OH)3,  together  with  Fe(OH)3. 
SALTS : 

CRYOLITE,  AlF3.3NaF. 
ALUM,  A12(SO4)3,  crystallized  with  a  sulphate  of  a  univalent     , 

metal. 

SILICATES,  particularly  the  feldspars  and  the  weathering 
products  of  the  feldspars,  e.g.  CaO.Al2O3.2SiO2. 

Extraction 

It  is  impossible  to  obtain  a  crude  Al  and  to  refine  it  by  the 
smelting  processes  such  as  have  been  described  for  other  metals, 
because  the  metal  is  too  electro-positive.  The  only  ores  that  come 
into  consideration,  therefore,  are  those  which  occur  in  a  relatively 
pure  state,  or  which  can  be  purified  without  difficulty.  Cryolite 
belongs  to  the  former  class  and  hydrargillite,  or  bauxite,  to  the 
latter.  In  the  present  method  of  preparing  aluminium,  cryolite  is 
used  as  solvent  for  bauxite. 
I.  Purification  of  Crude  A1(OH)3.  This  is  accomplished  by  the 

following  operations: 

i.  Transformation   into  Na3A103.      According  to  the  old-fash- 
ioned way  of  roasting  with  soda 

2A,(OH)3  +  3Na2CO3  =  2Na3AlO3  +  3H2O  +  3CO2, 

253 


254  METALLURGY 

or  by  dissolving  in  caustic  soda,  carrying  out  the  reaction 
in  iron  kettles  under  a  pressure  of  5  to  7  atmospheres. 


In  the  last  process  it  is  necessary  to  roast  the  ore  slightly 
before  treating  with  the  caustic  soda,  so  that  the  FeO  will 
be  converted  into  Fe2O3. 

2.  Leaching  of  the  Sodium  Aluminate  in  the  old  process;    in 
both  cases  there  is  filtration,  leaving  NaaAlOs  in  solution,  and 
a  residue  of  Fe2O3  and  the  gangue  of  the  ore. 

3.  Decomposition  of  the  Aluminate  by  C02. 


The  Al(OH)s  separates  out  from  the  solution  and  is  obtained 

by  filtration,  washing  and  drying. 

II.  Electrolysis  of  A1203  in  a  Molten  Bath.  The  Heroult  proc- 
ess is  based  upon  the  electrolytic  decomposition  of  A12O3  held 
in  solution  by  melted  AlF3.3NaF,  using,  at  the  same  time,  the 
electric  current  for  fusing  the  electrolyte  and  deposited  metal. 

The  working  conditions  as  carried  out  to-day  are: 

ELECTROLYZING  VESSELS:  wrought  iron  vats  with  rectangu- 
lar cross-section,  3  to  4.5  ft.  long,  20  to  30  in.  wide. 
The  vats  are  lined  with  carbon  at  the  bottom  and  on 
the  sides  with  cryolite.  Both  of  these  linings  are  kept 
cold  enough  by  air,  which  is  made  to  play  about  the  vessels 
by  artificial  draft,  so  that  no  O  enters  into  the  metal  from 
the  bottom,  and  none  of  the  side  lining  melts. 

ELECTROLYTE:   A12O3  dissolved  in  fused  AlF3.3NaF. 

ANODES:  Carbon  blocks  introduced  into  the  vat  from  above. 

CATHODES:  at  the  start,  the  carbon  bottom  lining  of  the 
vessel;  later,  the  metallic  Al  that  separates  out  there. 

TEMPERATURE:    750°  C.  (1382°  F.). 

CURRENT  DENSITY:  700  amperes  per  square  foot  of  cathode 
surface  (the  horizontal  cross-section  of  the  bath). 

BATH  POTENTIAL:  The  loss  in  E.M.F.  corresponds  on  an 
average  to  7.5  volts  per  furnace. 


ALUMINIUM 


255 


CONSUMPTION  OF  POWER:  1400  electric  horse  power  per  ton 
(2200  Ibs.)  of  metal  produced  in  24  hours. 

PROCEDURE  AND  REACTIONS  DURING  THE  ELECTROLYSIS, 
To  start  the  process,  the  upper  electrodes  are  lowered  until 
they  form  an  electric  arc  with  the  carbon  bottom,  then 
A1F3  is  charged,  and  the  electric  arc  is  maintained  until 
the  amount  of  fused  cryolite  is  sufficient  for  the  upper 
electrodes  to  dip  into  it,  then  the  addition  of  cryolite  is 


zzzzzzz^^ 

I 

FIG.  216. — Electric  Furnace  of  Heroult  Type. 

continued  until  the  melt  is  8  to  12  in.  deep  in  the  vat. 
Now,  as  Al  separates  out  during  the  electrolysis,  it  is 
replaced  in  the  bath  by  a  corresponding  amount  of  A12O3, 
carefully  avoiding  a  large  excess  of  the  latter.  The  metallic 
Al  separates  out  at  the  bottom  cathode  and  O  is  evolved 
at  the  anodes  which  dip  into  the  melt,  and  this  O  combines 
with  the  C  to  form  CC>2.  The  Al  is  removed  from  time 
to  time  by  tapping  or  ladling. 


256  METALLURGY 

Although  quite  a  number  of  other  electrolytic  processes  have 
proved  suitable  for  the  production  of  aluminium  on  a  large  scale, 
they  have  not  shown  so  great  economy  as  the  Heroult  process. 
They  possess  a  historical  interest,  however,  and  a  few  will  be 
mentioned  as  representing  the  gradual  development  of  the  Al 
industry. 

BUNSP:N  in  1854  deposited  the  first  Al  from  a  fused  compound, 
A1C13. 


FIG.  217. — Dendritic  Structure  on  Lower  Side  of  Cast  Ingot. 

ST.  CLAIRE  DEVILLE,  a  little  later  in  the  same  year,  proposed 
that  the  deposited  Al  be  replaced  in  the  bath  by  A12O3, 
but  soon  after  gave  up  electrolysis  and  proposed  the  decom- 
position of  AlCl3.NaCl  by  Na,  the  method  that  was  used  for 
a  long  time  in  the  technical  production  of  aluminium. 

ROSE,  in  1855,  recommended  the  decomposition  of  cryolite 
byNa. 

BEKETOFF,  in  1865,  the  decomposition  of  cryolite  by  Mg. 

GRABAU,  in  1887,  started  with  A12(SO4)3  and  discovered  the 
following  interesting  reactions: 

A12(SO4)3  +  2(AlF3.3NaF)  -4A1F3  +  3Na2SO4; 
4  A1F,; + 3Na2  =  A12  +  2  ( AlF^NaF) . 


ALUMINIUM  257 

COWLES  BROTHERS,  in  1884,  carried  through  the  reducing 
smelting  of  Al2Os  with  C  in  the  electric  furnace  but  could 
not  obtain  pure  Al;  they  worked,  therefore,  for  the  produc- 
tion of  Al  alloys  with  Cu  and  Fe:  aluminium  bronze  and 
ferro-aluminium. 
Properties  of  Aluminium : 

SPECIFIC  GRAVITY:   2.53. 

COLOR:   white  with  high  luster. 

MECHANICAL   PROPERTIES:     at    100°  to    150°  C.  (212°   to 
302°  F.),  it  can  be  forged  and  rolled  satisfactorily;  at  500°  C. 


FIG.  218. — Large-grained  Inner  Surface  of  a  Partly  Solidified  Block 
•    :  from  which  the  Still-molten  Metal  has  been  Withdrawn. 


(932°  F.)  it  can  be  compressed  easily;  at  530°  C.  (986°  F.) 
it  becomes  so  friable  that  it  can  be  pulverized. 

STRUCTURE:  according  to  the  way  it  is  poured,  it  is  either 
dendritic  or  grained. 

MELTING-POINT:   687°  C,  (1270°  F.). 

ALLOYS  with  most  other  metals.  In  the  experiments  carried 
out  in  the  attempt  to  prepare  Al,  and  as  regards  technical 
importance,  the  alloys  with  the  following  metals  have 
been  studied:  Cu,  Ni,  Co,  Fe,  Sb,  Sn,  Zn,  Mg. 

CHEMICAL  BEHAVIOR.  At  ordinary  temperatures  Al  is  very 
resistant  to  the  constituents  of  the  atmosphere;  even  at 
700°  to  800°  C.  it  oxidizes  only  slowly,  but  with  the  develop- 


258  METALLURGY 

ment  of  a  great  deal  of  heat  i  kg.  =  62  74  Cai.  (i  Ib.  = 
11,293  B.T.U.).  It  combines  with  halogens  even  at  low 
temperatures  and  readily  with  the  other  metalloids;  even 
with  N,  at  high  temperatures.  In  spite  of  its  high  solution 
tension  (  + 1.276  to  ward  H)  it  dissolves  comparatively  quickly 
only  in  HC1  and  NaOH;  in  H2SO4  it  dissolves  much  less 
readily,  and  in  HNOs  very  slowly.  Al,  however,  is  a  very 
energetic  reducing  agent  and  precipitant  for  metal  oxides 
and  other  compounds  of  metals,  both  in  the  solid,  molten 
and  dissolved  condition.  Particularly  in  the  decomposition 
or  oxides  rich  in  O,  so  much  heat  is  evolved  that  difficultly- 
fusible  metals,  like  Cr  and  the  slag  of  Al2Os,  are  melted 
or  even  volatilized  (cf.  Chromium,  page  222).  In  the  com- 
pounds that  are  of  industrial  importance,  Al  is  either 
present  as  a  trivalent  cation  (of  the  type  ALOs)  or,  when 
combined  with  strongly  basic  oxides,  in  the  anion  (forming 
aluminates,  of  the  type 


INDEX 


Acids,  washing  mercury  with,  59 

Allen  furnace,  67 

Allis-Chalmers  furnace,  copper,  76 

— ,  lead,  128 

Alloying  iron  in  the  bath,  215 
Aluminium,  253 
— ,  Beketoff,  256 
— ,  Bunsen,  256 
— ,  Cowles  Brothers,  257 
— ,  Deville,  256 
— ,  Grabau,  256 
— ,  HeYoult,  254 

—  hydroxide,  purification  of,  253 

—  oxide,  electrolysis  of,  257 
— ,  properties  of,  257 

— ,  Rose,  256 
—  salts,  253 
Amalgamation,  gold,  2 
— ,  silver,  36 

— ,  — ,  caldron  process,  46 
— ,  — ,  Caso  process,  46 
— ,  — ,  Krohnke  process,  41 
— ,  — ,  Patio  process,  41 
— ,  — ,  Washoe  process,  46 
— ,  — ,  with  chemicals,  40 
— ,  — ,  without  chemicals,  36 
Amalgamator,  Laszlo,  40 
— ,  Schemnitz,  40 
Amalgam  catchers,  38 
Anaconda  copper  converters,  97, 100 
Antimonial  lead,  154 
Antimony,  149 

— ,  Borchers'  furnaces,  112,  114 
— ,  concentration,  149 


Antimony,  electrolysis  152 
— ,  properties,  156 
— ,  refining,  155 
— ,  sources,  149 
— ,  starring,  155 
—  ores,  liquation,  149 
— ,  lixiviation,  151 
,  oxidizing  roast  with  sublima- 
tion, 150 

,  precipitation  methods,  152 

,  reduction  methods,  151 

Argall  furnace,  copper,  70,  71 
Arsenic,  removal  from  tin  ores,  134 
Augustin  process,  silver,  48 
Austen,  see  Roberts- Austen 
Austenite,  192 

B 

Barrel  amalgamation,  silver,  47 
Base  bullion,  29 
Bath,  alloying  iron  in  the,  215 
Beketoff,  256 

Belgian  treatment  of  zinc  ores,  236 
Benedicks,  constitution  of  iron-car- 
bon alloys,  192 
Bertrand-Thiel  process,  203 
Bessemer  process,  202 
Betts,  electrolysis  of  lead,  130 
Bismuth,  109 

— ,  Borchers'  furnaces,  112,  114 
— ,  electrolysis,  116 
— ,  melting  with  sulphur,  116 
— ,  oxidizing  fusion,  115 
— ,  precipitation,  115 
— ,  properties,  116 

259 


260 


INDEX 


Bismuth  ores,  109 

,  chemical  concentration,  109 

,  concentration  processes,  109 

,  liquation,  110 

,  melting  in  crucibles,  111 

,  melting  in  reverberatories,  114 

,  reduction  process,  110 

Blast  furnaces,  iron,  181,  182,  183 
— ,  — ,  reactions  in,  185 
— ,  lead,  128 
Blast,  hot,  182 

Blowing  copper  matte,  Manh&s,  95 
—  hi  converters,  iron,  201 

— ,  nickel,  162 

Bottoms,  smelting  of,  nickel,  165 
Borchers,   continuous   kiln,  bismuth 

and  antimony,  112 
— ,  crucible  furnace  with  regenerative 

chambers,  chromium,  219 
— ,  direct  electric  treatment  of  nickel 

matte,  173 
— ,  electric  smelting  of  nickel  matte, 

171 

— ,  electric  smelting  of  tin,  138 
— ,  electrolytic  treatment  of  copper 

matte,  106 
— ,  reverberatory    furnace,    bismuth 

and  antimony,  114 
— ,  treatment  of  nickel  ores,  168 
Bradford,  converter  for  lead  ores,  127 
Brittany  process,  lead,  121 
Broken  Hill,  zinc  furnace,  247 
Brown  furnaces,  copper,  67,  68 
Bruckner  furnace,  copper,  69,  70 
Bunsen,  256 
Butters'  vacuum  filter,  gold,  19 


Cadmium,  229 
—  ores,  229 
— ,  properties  of,  230 
Caldron  process,  silver,  46 
Capacity  of  electric  furnaces,  213 
Carbon-iron   alloys,   constitution   of, 
188,  192 

— ,  Benedicks,  192 
,  Charpy,  192 


Carbon-iron  alloys,  Goerens,  188,  192' 
,  Roberts-Austen,  188,  192 

— ,  Roozeboom,  192 

,  Wiist,  188,  192 

Carinthian  process,  lead,  121 
Carmichael-Bradford  converter,  lead,, 

126,  127 

Case-hardening  of  steel,  216 
Caso  process,  silver,  46 
Cast  iron,  180 

— ,  gray,  191 

—  — ,  white,  190 
Cementation,  216 
Cementite,  188,  192 
Charcoal  hearth  process,  199 
Charpy,  constitution  of  iron-carbon 

alloys,  192 

Chemical  concentration,  bismuth,  109 
— ,  tin  ores,  133 

—  solution  and  precipitation,  gold,  2 

,  platinum,  26 

Chloridizing  fusion,  silver,  51 

—  roasting,  muffles  for,  72 
Chlorinating  solution,  gold,  21 
Chlorination  barrel,  gold,  4 

-  gold,  2 

—  plant,  5,  6 

—  platinum,  26 

Chromic  oxide,  reduction  of,  222 
Chromite,  reducing  smelt  of,  218 
Chromium,  218 

—  ores,  218 

— ,  properties  of,  223 

— ,  separation  from  iron,  220 

—  solutions,  electrolysis  of,  222 
Colby  furnace,  iron,  211 

Colorado  Iron  Works,  furnaces,  76,  77 
Concentrated   nickel  matte,   electric 

smelting  of,  174 
Concentrated     nickel     matte,     pure 

nickel  from,  173 

Concentration,  of  antimony  ores,  149 
— ,  of  bismuth  ores,  109 
— ,  of  copper  ores,  60,  90 
— ,  of  iron  ores,  178 
— ,  of  lead  ores,  118 
— ,  of  nickel  ores,  160 


INDEX 


261 


Concentration,  of  tin  ores,  133 
Converters,  Anaconda,  97,  100 
— ,  blowing  in,  nickel,  162 

—  for  copper  ores,  97,  100 
—  for  iron  ores,  203,  204,  205 

—  for  lead  ores,  126,  127 

— ,  Carmichael-Bradford,  126,  127 

— ,  Huntington-Heberlein,  126 

— ,  Savelsberg,  127 

— ,  Stalmann,  96,  98,  99 

— ,  Sticht,  96,  98,  99 

Converting  copper  matte,  89 

Constitution   of    iron-carbon    alloys, 

188-193 
Copper,  60 
— ,  Allen  furnace,  67 
— ,  Allis-Chalmers  furnace,  76 
— ,  Anaconda  converters,  97,  100 
— ,  Argall  furnace,  70,  71 
— ,  Brown  furnace,  67 
— ,  Bruckner  furnace,  70 
— ,  Colorado  Iron  Works  furnaces,  76, 

77 

—  concentrating  processes,  60 

—  converters,  96-99 

—  enriching  processes,  60 

—  fore-hearths,  81 

— ,  furnace  refining,  102 

— ,  hand  reverberatories,  66 

— ,  heap  roasting,  63 

— ,  hearth  furnaces,  66 

— ,  Herreshoff  furnace,  66,  67 

— ,  Hixon  furnace,  67 

— ,  Hocking-Oxland  furnace,  70 

— ,  Holthoff-Wethey  furnace,  67 

— ,  Humboldt  furnace,  67 

— ,  Johnson  furnace,  80 

— ,  Kaufmann  furnace,  67 

— ,  Keller  furnace,  69 

— ,  kernel  roasting,  87 

—  leaching,  concentration  by,  90 
— ,  lump  pyrite  furnace,  Lunge,  64 
— ,  MacDougall  furnace,  66,  67 

— ,  Maletra-Schaffner  furnace,  64 

— ,  Mansfeld  furnaces,  85 

— ,  —  kiln,  64 

— ,  Mathewson  furnaces,  83,  87 


Copper     matte,     concentration    by 

smelting,  89 

,  converting,  89,  95 

,  electrolytic  treatment,  106 

,  nature  of,  Rontgen,  89 

— ,  oxidizing  roast  of,  89 
— ,  reverberatory  furnaces  for,  84 

— ,  shaft  furnaces,  73 

— ,  smelting  for,  71,  87 
— ,  muffle  furnaces,  71,  72 
— ,  O'Hara  furnace,  67 
— ,  O'Hara-Brown  furnace,  68 
— ,  Oker  furnace,  75 

—  ores,  60 

— ,  oxidizing  roast,  61 
— ,  Parkes  furnace,  66 
— ,  Pearce  furnace,  67 
— ,  Peters  furnace,  86 
— ,  precipitation  of,  99 
— ,  properties  of  refined,  107 
— ,  purification    of    leached    liquor, 
Hofmann,  94 

—  pyrite  furnace,  63 
— ,  pyritic  smelting,  87 
— ,  reaction  smelting,  95 

— ,  reverberatories  as  fore-hearths,  81 
— ,  reverberatory  furnaces,  65,  84 
— ,  roasting,  65 

—  — ,  smelting,  89 

—  roasting  furnaces,  66-73 

—  —  operations,  61 

—  roast-reduction  process,  99 

—  slag  and  matte  spouts,  81 

—  calculation,  73 

—  smelting  for  matte,  71,  87 

—  operations,  61 

— ,  solution  of  gold  in,  1 
— ,  of  platinum  in,  25 

—  — ,  of  silver  in,  28 
— ,  Spence  furnace,  69 
— ,  stall  roasting,  63 

— ,  Stalmann  converter,  96,  98,  99 

— ,  Stetefeldt  furnace,  65 

— ,  Sticht  converter,  96,  98,  99 

— ,  —  furnace,  78,  79 

— ,  Walker  casting  machine,  103 

— ,  Wethey,  furnace,  69 


262 


INDEX 


Copper,  White-Howell  furnace,  70 
Cowles  Brothers,  aluminium,  257 
Cowper  stoves  for  heating  blast,  183 
Crucible  furnace,  chromium,  219 
Crucibles,  melting  in,  bismuth,  111 
Crucible  steel,  215 
Crude  nickel  matte  or  speiss,  162 
Crystallization  of  tin,  144 
-  processes,  30,  130 
Cupellation,  English  and  German 

methods,  33 

Cupelling  furnaces,  37,  38 
Cyanide  leaching,  7 

—  plant  for  gold  ores,  12 

—  solutions,  electrolysis  of,  gold,    18 
— ,  precipitation  of  gold  from,  15 

D 

Darby-Phoenix  process  for  recarbur- 

izing  iron,  215 
De   Ferranti,  induction   furnace  for 

iron  smelting,  211 
Distillation  of  mercury,  59 

—  of  zinc,  245 
Deville,  St.  Claire,  256 

Direct     electrolytic      treatment     of 

nickel  matte,  173 
Dotsch,  copper  leaching,  94 
Douglas,  copper  leaching,  94 
Desulphurizing  fluxes,  heating  with, 

mercury,  59 
Dross,  118 

E 

Electric  furnace,  aluminium,  255 
— ,  iron  and  steel,  211-213 

—  smelting,  nickel  matte,  174 
— ,  steel,  207 

,  tin,  138 

Electrolysis,  aluminium  oxide,  254 

— ,  antimony,  152,  156 

— ,  bismuth,  116 

— ,  chromium  solutions,  222 

— ,  copper,  104 

— ,  matte,  106 
-,  gold,  22 
— ,  lead,  130 
— ,  nickel,  172 


Electrolysis,  silver,  53 

— ,  tin,  141 

— ,  zinc,  246 

English  method   of  cupelling  silver 

bullion,  33 

—  roast-reaction  process,  lead,  121 
Extraction,  aluminium,  253 
— ,  antimony,  151 
— ,  bismuth,  110 
— ,  cadmium,  229 
— ,  copper,  95 
— ,  gold,  1 
— ,  lead,  119 
— ,  mercury,  56 
— ,  nickel,  169 
— ,  platinum,  25 
— ,  silver,  29 
— ,  tin,  134 
— ,  zinc,  232 


Faber  du  Faur's  furnace,  zinc,  247 

Ferrite,  192 

Ferro-chrome,  218 

Ferro-manganese,  251 

Ferro-molybdenum,  196 

Ferro-silicon,  196 

Ferro-tungsten,  225 

Ferro- vanadium,  196 

Filter  press,    Klein,   Schanzlin   and 

Becker,  14 
— ,  London  and    Hamburg  Gold 

Recovery  Co.,  17 
Filter  tank  for  gold,  10 
Filtration  of  mercury,  59 
Fire  refining  of  crude  silver,  53 
Fore-hearths,  copper,  81 
Franke,    electrolytic    treatment    of 

copper  matte,  106 
Freiberg  vitriolization  process,  51 
French  of  Brittany  process,  lead,  121 
Fluxes,  desulphurizing  for  mercury, 59 
Forgeable  iron,  197 

— ,  finishing  of,  215 
Furnace  gas,  iron,  194 
Furnaces,  Allen,  for  copper  ores,  67 
— ,  Allis-Chalmers  for  copper  ores,  76 


INDEX 


263 


Furnaces,    Allis-Chalmers     for    lead 

ores,  128 

— ,  American  water-jacket,  77 
— ,  Argall  for  copper  ores,  70 
— ,  Borchers  and  Mattonet,  tin,  138 
—  — ,  antimony  and  bismuth,  112, 114 
— ,  —  crucible,  chromium,  219 
— ,  blast,  lead,  127 
— ,  — ,  iron,  183 
— ,  Brown,  copper,  67 
— ,  Bruckner,  69,  70 
— ,  capacity  of  electric,  213 
— ,  Chinese,  for  tin  ore,  136,  137 
— ,  Colby  induction,  iron,  211 
— ,  Colorado  Iron  Works,  copper,  76, 

77 

— ,  De  Ferranti  induction,  iron,  211 
— ,  Faber  du  Faur,  zinc,  247 
— ,  Girod;  electric  for  iron,  209,  210 
— ,  Halbhochofen,  lead,  127,  128 
— ,  hand      reverberatories,      copper, 

66 

— ,  hearth  reverberatories,  66 
— ,  Heroult,  aluminium,  255 
— ,  — ,  iron,  208 
— ,  Herreshoff,  copper,  66,  67 
— ,  Hixon,  copper,  67 
— ,  hochofen,  lead,  127,  128 
— ,  Hocking-Oxland,  copper,  70 
— ,  Holthoff-Wethey,  67 
— ,  Humboldt,  copper,  67 
— ,  Johnson,  copper,  80 
— ,  Kaufmann,  copper,  67 
— ,  Keller,  copper,  69 
— ,  Kjellin,  iron,  211 
— ,  Krupp,  for  amalgam,  46 
— ,  Krohnke,  for  distilling  amalgam, 

'46 

— ,  kriimmofen,  lead,  127 
— ,  lump  pyrite,  Lunge,  64 
— ,  MacDougall,  copper,  66,  67 
— ,  Maletra-Schaffner,  copper,  64 
— ,  Mansfeld,  copper,  64,  85 
— ,  Mathewson,  copper,  83,  84,  87 
— ,  Mattonet,  tin,  138 
— ,  muffle,  for    chloridizing    copper 
ores,  71,  72 


Furnaces,  muffle,  for  gold  ores,  8 
— ,  North     American     water-jacket, 

lead,  129 

— ,  O'Hara,  copper,  67 
— ,  O'Hara-Brown,  copper,  68 
— ,  Oker,  copper,  75 
— ,  Parkes,  copper,  66 
— ,  Pearce,  copper,  67 
-  -,  gold,  3 
— ,  Peters,  copper,  86 
— ,  reverberatories  as  fore-hearths,  81 
— ,  — for  bismuth  and  antimony  ores, 

114 

— ,  —  for  copper  matte,  84 
— ,  -*—  for  copper  ores,  65 
— ,  —  for  lead,  125 
— ,  —  for  oxidation  of  impurities  in 

iron,  205 

— ,  —  with  moving  hearths,  69 
— ,  Rochling-Rodenhauser,  212,  213 
— ,  Schrubko,  zinc,  237,  238 
— ,  shaft,  copper,  73 
— ,  — ,  lead,  127 

— ,  Siemens  open  hearth,  iron,  206 
— ,  Spence,  copper,  69 
— ,  Stetefeldt,  copper,  65 
— ,  Sticht,  copper,  78,  79 
— ,  Upper  Harz,  lead,  128 
— ,  Wethey,  copper,  69 
— ,  White-Howell,  copper,  70 
Fusion,  chloridizing,  of  silver,  51 
— ,  desulphurizing    of    nickel  matte, 

171 

— ,  liquating  of  zinc,  245 
— ,  oxidizing  of  bismuth,  115 
— ,  —  of  tin,  144 
— ,  reducing  of  tin  ores,  135 
— ,  —  of  wolframite  or  scheelite,  225 
— ,  sulphurizing  of  antimony,  155 
— ,  —  of  bismuth,  116 
— ,  —  of  silver,  51 


German  method  of  cupelling  silver 

bullion,  33 
Girod  furnace,  iron,  239,  210 


264 


INDEX 


Goerens,  constitution  of  iron-carbon 

alloys,  188,  192 
Gold,  1 

—  amalgamation,  2 

— ,  American  chlorination  barrel,  4 
— ,  Butters  vacuum  filter  for  slimes,  19 
— ,  chemical  solution  and  precipita- 
tion, 2 

—  chlorinating  solution,  21 

—  chlorination,  2 
barrel,  4 

—  plant,  5,  6 

—  cyanide    leaching   and    precipita- 
tion, 7 

-  plant,  12 

—  electrolysis,  18 

—  electrolytic  parting,  22 

—  extraction,  1 

—  filter  press,  14 

—  plant,  17 

-  vat,  10 

—  leaching  methods,  recent,  11 

—  muffle  furnace,  8 

—  ore  dressing,  1 

—  ores,  1 

-  parting,  20,  49 
— ,  Pearce  turret  furnace,  3 

—  precipitation  from  cyanide  solu- 
tions, 15 

—  by  zinc,  16 

—  refining,  21 

—  slimes,  Butters  filter  for,  19 

—  solution  and  precipitation,  21 

—  in  metals,  1,  2 

— ,  —  of  platinum  in,  25 

—  wet-tube  mill,  Krupp,  13 
Goldschmidt,    process    for    tin-plate 

waste,  141 

— ,  thermite  process  for  chromium,222 
— ,  —     —  for  manganese,  252 
Grabau,  aluminium,  256 
Graphite,  in  iron,  193 
Gray  cast  iron,  191 
Giinther,    electrolytic    treatment    of 

copper  matte,  106 
— ,  pure    nickel    from    concentrated 

matte,  173 


H 

Halbhochofen,  lead  ores,  127,  128 
Hand  reverberatories,  copper,  66 
Hard  lead,  154 

Harz  vitriolization  process,  silver,  53 
Heap  amalgamation,  silver,  40 

—  roasting  of  copper  ores,  63 

—  of  lead  ores,  124 
Hearth  furnaces  for  copper  ores,  66 

—  process  for  lead  ores,  121 
Hearths,  reverberatories  with  mov- 
ing,, copper,  69 

He"roult,  aluminium,  254,  255 

—  furnace  for  iron  and  steel,  208 
Herrenschmidt's   process   for   nickel 

ores,  167 

Herreshoff  furnace  for  copper  ores,  67 
Hixon  furnace  for  copper  ores,  67 
Hocking-Oxland  furnace  for   copper 

ores,  70 

Hofmann,  copper  leaching,  94 
—  process  for  treating  sliver  ores,  49 
Holthoff-Wethey  furnace  for  copper 

ores,  67 
Hopfner,  lixiviation  and  electrolytic 

deposition,  zinc,  246 
Hot  blast,  iron,  182 
Hunt  and  Douglas,  copper  leaching, 

94 
Huntington-Heberlein  converter  for 

lead  ores,  126 


Iron,  177 

—  alloying  in  the  bath,  215 

—  alloys,  196 

—  annealing  steel,  198 

— ,  Basic  Bessemer  process,  202 
— ,  Bertrand-Thiel  process,  203 
— ,  Bessemer  process,  202 
— ,  blast  furnaces,  181 
— ,  —  — ,  reactions  in,  185 
— ,  capacity  of  electric  furnaces;  213 
— carbon    alloys,    constitution    of, 

188,  192 
— ,  case  hardening,  216 

—  cementation,  216 


INDEX 


265 


Iron  charcoal-hearth  process,  199 

—  Colby  induction  furnace,  211 

—  compounds,  196 

—  converters,  203 

— ,  Cowper  stoves  for  blast,  183 

— ,  de  Ferranti    induction    furnace, 

211 

— ,  electric  furnaces,  208-213 
— ,  —  smelting,  207 
— ,  forgeable  197 
— ,  — ,  finishing  of,  215 

—  furnace  gas,  194 

— ,  Girod  furnace,  209,  210 

— ,  hardening  of  steel,  197 

— ,  HeVoult  furnace,  208 

— ,  induction  furnaces,  211 

— ,  kish,  189 

— ,  Kjellin  induction  furnace,  211 

— ,  malleableizing,  198 

— ,  mixed  crystals,  189 

— ,  open-hearth  process,  203 

—  ores,  177 

— ,  concentrating  and  other  pre- 
liminary operations,  178 
— ,  reducing  roast,  180 

— ,  smelt,  180 

— ,  oxidizing  fusion  of  cast,  201 
-,  pig,  188 

—  products  of  the  blast  furnace,  188 
— ,  properties,  216 

-  puddling,  200 

—  recarburization,  (Darby-Phoenix) 
215 

— ,  Rochling  -  Rodenhauser    furnace, 

212,  213 

— ,  separation  from  chromium,  220 
— ,  —  from  tungsten,  226 
— ,  Siemens  open-hearth  furnace,  206 
— ,  Siemens-Martin  process,  203 

-  slag,  193 

— ,  Talbot  process,  203 

— ,  Thomas-Gilchrist  process,  202 

—  welding,  215 

— ,  Wittkowitz  process,  207 


Johnson  furnace,  copper,  80 


K 

Kaufmann  furnace,  copper,  67 

Keller  furnace  copper,  69 

Kernel  roasting,  87 

Kiln,  Mansfeld  for  copper  ores,  64 

Kish,  189 

Kiss  process,  48 

Kjellin  induction  furnace,  211 

Klein,  Schanzlin  and  Becker,  filter 
press,  14 

Krohnke  furnace  for  distilling  amal- 
gam, 46 

—  process  of  silver  amalgamation,  41 
Krummofen,  127 

Krupp  furnace  for  distilling  amalgam, 
46 

—  wet  tube  mill  for  gold  ores,  13 


Laszlo  amalgamator,  silver,  39 

Leached  liquor,  purification  of,  cop- 
per, 94 

Leaching,  concentration  of  antimony 
ores,  by,  151 

— ,  —  of   copper  ores  by,  90 

— .  —  of  nickel  ores  by,  166 

— ,  —  of  zinc  ores  by,  246 

Lead,  118 

— ,  Allis-Chalmers  furnace,  128 

— ,  Arent  tap  furnace,  129 

— ,  Brittany  process,  121 

— ,  Carinthian  process,  121 

— ,  Carmichael-Bradford  converter, 
126,  127 

— ,  —  treatment  in  a  converter,  122 

— ,  concentration  of  ores,  118 

—  electrolysis,  Betts,  130 
— ,  English  process,  121 
— ,  French  process,  121 

—  halbhochofen,  128 
— ,  heap  roasting,  124 
— ,  hearth  process,  121 

— ,  Huntington-Heberlein  converter, 

126 

— ,  —  method,  122 
— ,  krummofen,  127 


266 


INDEX 


Lead,  liquation  and  crystallization 
processes,  130 

—  matte,  124 

— ,  North- American  water-jacket  fur- 
nace, 129 

—  ores,  118 

— ,  concentration  of,  118 
— ,  oxidation  of,  130 
— ,  Pilz-Freiberg  furnace,  128 

—  precipitation  process,  123 
— ,  properties  of,  130 

—  reaction  smelting,  119 

—  reduction  smelting,  123 

— ,  reverberatory  furnaces,  125 

-  roasting,  119,  122 

—  roast  reduction  process,  122 
— ,  Savelsberg  converter,  127 
— ,  Schenck  and  Rassbach,  119 
— ,  shaft  furnaces,  127 

— ,  sinter  roasting,  125 

— ,  slag  roasting,  125 

— ,  solution  of  gold  in,  2 

— ,  —  of  platinum  in,  25 

- — ,  —  of  silver  in,  28 

— ,  Tarnowitz  process,  121 

Lehmer's  process  for  treating  nickel 

matte,  171 
Liquation  of  antimony,  149 

—  of  bismuth,  110 

-  of  tin,  143 

—  of  zinc,  245 

London  and  Hamburg  Gold  Recovery 
Co.  filter  pressing  plant,  17 

Luce-Rozan  process,  silver  extraction, 
30 

Lunge  lump  pyrite  furnace,  copper,  64 

M 

MacDougall  furnace  for  copper  ores, 

66,  67 
Maletra-Schaffner  furnace  for  copper 

ores,  64 

Malleableizing,  198 
Manganese,  250 

—  alloys,  250 

—  compounds,  251 

— ,  Goldschmidt  thermite  process,252 


Manganese  ores,  250 

— ,  properties,  252 

— ,  separation  from  tungsten,  226 

—  silicides,  251 

Manhes  converter  for  copper  matte,96 
Mansfeld  furnace  for  copper  ores,  85 

—  kiln  for  copper  matte,  64 
Martensite,  192 

Mathewson  furnaces  for  copper  ores, 

83,  84,  87 
Matte,     copper,      concentration     by 

smelting,  89 

— ,  — ,  converting,  89,  95 
— ,  — ,  furnaces  for,  84 
— ,  — ,  oxidizing  roast  of,  89 
— ,  — ,  smelting  for,  71,  87 
— ,  nickel,  concentration  of,  162 
— ,  — ,  converting  of,  162 
— ,  — ,  formation  of,  162,  164 
— ,  — ,  oxidizing  roast  of,  162 

—  spouts,  81 

Mattonet,  smelting  of  tin  in  an  electric 

furnace,  138 
Mechanical  concentration  of  tin  ores, 

133 
Mehler,  press  for  making  zinc  retorts, 

243 

Melting  bismuth  ores  in  crucibles,  111 
—  in    reverberatory    furnaces, 

114 
Mercury,  56 

—  distillation,  59 

—  filtration,  59 

—  furnaces,  57,  58 

— ,  heating    with    desulphurizing 
fluxes,  59 

—  ores,  56 

— ,  roasting  of,  56 
— ,  oxidizing  roast,  56 
— ,  washing  with  acids,  59 
Metals,  solution  of  gold  in,  1 
— ,  —  of  platinum  in,  25 
— ,  —  of  silver  in,  28 
Mixed  crystals  in  iron-carbon  alloys, 

189 
Mond's  process  for  extracting  nickel, 

174 


INDEX 


267 


Mortar  amalgamation,  silver,  39 
Muffle  furnaces,' for  chloridizing  roast 
of  copper  ores,  72 

,  for  copper  ores,  71 

— ,  for  gold  ores,  8 
Muffles,  types  of,  for  zinc  ores,  236 

N 

Nickel,  160 

— ,  blowing  in  converters,  162 

— ,  Borchers  and  Warlimont's  proc- 
ess, 168 

— ,  concentration  and  elimination  of 
copper,  164 

,  by  leaching  and  precipitation, 

166 

,  of  the  matte  or  speiss,  162 

,  without  separation  of  copper, 

161 

— ,  desulphurizing  fusion  of  concen- 
trated matte,  171 

—  electrolysis,  172 

—  extraction,  169 

— ,  formation  of  first  matte  or  speiss, 

162 
— ,  Herrenschmidt's  process,  167 

—  matte,  electric  smelting  of,  174 

—  ores,  160 

,  concentration     methods,     160 

— ,  oxidizing  roast,  161,  164 

— ,  —    —  of  matte  or  speiss,  162 

— ,  properties,  175 

— ,  refining  fusion,  172 

— ,  roast  reaction  smelting,  170 

— ,  —  reduction  process,  169 

— ,  smelting  of  the  bottoms,  165 

— ,  —  of  the  tops,  165' 

— ,  —  to  a  matte,  164 

— ,  solution  of  platinum  in,  25 

Nitric  acid  parting  of  silver  and  gold, 

49 
Non-sinter  roasting  of  lead  ores,  126 


O'Hara  furnace,  copper,  67 
O'Hara-Brown  furnace,  copper,  68 
Oker  process  of  silver  leaching,  48 


Oxidation  of  impurities  in  lead,  130 

-  of  lead,  130 

Oxidizing  fusion  of  bismuth,  115 

—  of  tin,  144 

—  roast  of  antimony  ores,  150 

—  of  copper  ores,  61 

—  of  mercury  ores,  56 

—  of  nickel  matte  or  speiss,  162 

—  of  nickel  ores,  161,  164 


Pan  amalgamation,  silver,  40,  47 

Parkes  furnace,  copper,  66 

—  process,  silver,  30 

Parting  gold  and  silver,  20,  49,  51 

Patera  process,  silver,  49 

Patio  amalgamation  process,   silver, 

40 

Pattinson  process,  silver,  30 
Pearce  turret,  furnace,  gold,  3 

—  roasting  furnace,  copper,  67 
Pearlite,  193 

Peters,  American   reverberatory  fur- 
nace, copper,  86 

Phoenix  recarburization  process,  iron, 
215 

Pig  iron,  188 
-  mixer,  203 

Pilz-Freiberg  furnace  for   lead  ores, 
128 

Platinum,  25 

— ,  chemical  solution  and   precipita- 
tion, 26 

—  chlorination,  26 

—  ores,  25 

— ,  properties,  27 

—  solution  in  metals,  25 
Poor  tin  slags,  smelting  of,  141 
Precipitation,  chemical  solution  and, 

of  gold,  2 

— ,  -         — ,of  platinum,  25,  26 
— ,  concentration  of  nickel  by,  166 

—  of  copper,  99 
-of  gold,  21 

—  methods,  antimony,  152 

— ,  bismuth,  115 
,  lead,  123 


268 


INDEX 


Precipitation    methods,   manganese, 
252 

— ,  zinc,  244 

Preliminary  operations,  iron  ores,  178 
Properties  of  aluminium,  257 

—  of  antimony,  156 

—  of  bismuth,  116 

—  of  cadmium,  230 

—  of  chromium,  223 

—  of  copper,  107 
—  of  gold,  23 

—  of  iron,  216 

-  of  lead,  130 

— -  of  manganese,  252 

—  of  mercury,  59 

-  of  nickel,  175 

—  of  platinum,  27 

—  of  silver,  54 

-  of  tin,  145 

—  of  tungsten,  227 

—  of  zinc,  248 
Puddling,  200 

Pyrite  furnace,  Lunge,  64 
Pyritic  smelting,  87 

R 

Rassbach,  reaction  smelting  for  lead, 
119 

Reactions  in  the  blast  furnace,  185 

Reaction  smelting,  copper,  95 
— ,  lead,  119 

Recarburization     of     iron,     Darby- 
Phoenix  process,  215 

Reducing  roast  of  iron  ores,  180 

—  smelt  of  chrome  iron  ore,  218 
of  iron  ores,  180 

—  of  tin  ores,  135 

—  of  tungsten  ores,  225 
Reduction  methods,  antimony  ores, 

151 

—  for  chromic  oxide,  222 

—  for  iron,  214 

—  for  tungstic  acid,  227 

—  process,  bismuth,  110 
Refining  antimony,  155 

—  base  bullion,  29 

—  bismuth,  115 


Refining  cadmium,  230 
—  copper,  101 

-  gold,  21 

—  lead,  129 

—  manganese,  252 

—  mercury,  59 

-  nickel,  171 

—  platinum,  26 

—  silver,  53 

-  tin,  143 

-  tungsten,  226 

—  zinc,  245 

Regenerative  firing,  chromium,  219 
— ,  steel,  203 

Reverberatory     furnaces,     as     fore- 
hearths,  copper,  81 

—  for  bismuth  ores,  114 

—  for  copper  smelting,  89 

—  for  copper  matte  smelting,  84 

—  for  copper  ores,  roasting  of,  65 

—  for  lead  ores,  125 

—  for  tin  ores,  136 

—  with  moving  hearths,  69,  126 
Rhenish  zinc  smelting,  239 

Rich  tin  slags,  smelting  of,  140 

Rio  Tin  to  leaching  plant,  copper  ores, 

93 
Roasting  antimony  ores,  150 

—  copper  ores,  61 

—  furnaces  for  copper  ores,  Allen,  67 

— ,  Argall,  70,  71 

— ,  Brown,  67 

— ,  Bruckner,  70 

— ,  Herreshoff,  66,  67 

— ,  Hixon,  67 

— ,  Hocking-Oxland,    70 

— ,  Holthoff-Wethey,  67 

— ,  Humboldt,  67 

— ,  Kaufmann,  67 

— ,  Keller,  69 

— ,  MacDougall,  66,  67 

— ,  O'Hara,  67 

— ,  O'Hara-Brown,  68 

— ,  Parkes,  66 

— ,  Pearce,  67 

— ,  Spence,  69 

— ,  Wethey,  69 


INDEX 


269 


Roasting  furnaces  for  copper  ores, 
White-Howell,  70 
— ,  for  gold  ores,  Pearce,  3 

-  lead  ores,  119,  125 

— ,  processes,  Brittany,  121 
— ,  — ,  Carinthian,  121 
— ,  — ,  English,  121 
— ,  — ,  Hearth,  121 
— ,  — ,  Tarnowitz,  121 

-  iron  ores,  178,  180 
—  mercury  ores,  56 

—  with  desulphurizing  fluxes, 
59 

—  muffles    for     chloridizing     burnt 
pyrite,  72 

—  nickel  ores,  161,  164 

—  matte  or  speiss,  162 
Roast-reaction  smelting,  nickel,  170 
Roast-reduction  process,  copper,  99 
— ,  lead,  122 
— ,  nickel,  169 
— ,  zinc,  232 

Roberts-Austen,  constitution  of  iron- 
carbon  alloys,  188,  192 
Rochling-Rodenhauser    electric    fur- 
nace for  iron  smelting,  212,  213 
Rontgen,  nature  of  copper  matte,  89 
Roozeboom,    constitution    of    iron- 
carbon  alloys,  192 
Rose,  aluminium,  256 
Rossler's  sulphurizing,  fusion,  silver, 

51 

Russell   process   for   leaching   silver 
ores,  49 

S 

Savelsberg  converter  for  lead  ores,  127 
Scheelite,  reducing  fusion  of,  225 
Schenck,  reaction  smelting  for  lead, 

119 

Schrubko,  zinc  furnace,  237,  238 
Separation,  electrolytic,  of  zinc,  246 
— ,  iron  from  chromium,  220 
— ,  tungsten  from  iron,  manganese, 

and  calcium,  226 
Shaft  furnaces  for  copper  and  lead, 

66,  73,  127 
,  Allis-Chalmers,  76,  128 


Shaft     furnaces,     American    water- 
jacket,  77,  80,  129 

— ,  Colorado  Iron  Works,  76,  77 

— ,  early  type,  75 

— ,  Johnson,  80 

— ,  Mathewson,  83 

— ,  Oker,  75 

— ,  Pilz,  128 

— ,  Sticht,  78,  79 

— ,  Upper  Harz,  128 
Siemens  open-hearth  furnace,   iron, 

206 

Siemens-Martin  process,  203 
Silesian  method  of  zinc  extraction, 

235 

Silver,  28 
— ,  amalgamation,  36 

— ,  apparatus,  47 

— ,  Caso  or  caldron  process,  46 

— ,  Krohnke  process,  41 

— ,  Patio  process,  40 

—  with  chemicals,  40 

—  without  chemicals,  36 

—  amalgam  catchers  and  auxiliary 
amalgamators,  38 

—  amalgam  filter,  45 

—  American  amalgamating  pan,  45 

—  arrastra,  40 

— ,  Augustin  process,  48 

—  barrel  amalgamation,  47 

—  chloridizing  fusion,  51 

—  concentration  of  ores,  28 

—  extraction,  28 

— ,  English  cupellation,  33 
— ,  —  cupelling  furnace,  38 

—  electrolysis,  51 

—  electrolytic  refining,  53 

— ,  Freiberg  vitriolization  process,  51 

— ,  German  cupellation,  33 

— ,  —  cupelling  furnace,  37 

— ,  Harz  vitriolization  process,  53 

— ,  Kiss  process,  48 

— ,  Krohnke    furnace    for    distilling 

amalgam,  46 
— ,     Krupp    furnace,    for    distilling 

amalgam,  46 
— ,  Laszlo  amalgamator,  39 


270 


INDEX 


Silver,  Luce-Rozan  process,  30 

—  mortar  amalgamation,  39 
— ,  nitric  acid  parting,  49 

— ,  Oker-Longmaid-Henderson  proc- 
ess, 48 
-  ores,  28 

—  pan  amalgamation,  40,  47 
— ,  Parkes  process,  30 

— ,  Patera  process,  49 
— ,  Pattinson  process,  30 

—  properties,  54 

— ,  refining  of  the  base  bullion,  29 
— ,  Rossler's  sulphurizing  fusion,  51 
— ,  Russell  process,  49 
—  skimmings  from  base  bullion,  29 

—  sluice  amalgamation,  38 

—  solution  in  metals,  28 
— ,  solution  of  gold  in,  2 

—  stamp  mill  amalgamation,  38 
— ,  sulphuric  acid  parting,  51 
— ,  vat  amalgamation,  47 

— ,  Washoe  process,  46 
— ,  Ziervogel  process,  47 
Sinter  roasting  of  lead  ores,  125 
Slag  and  matte  spouts,  copper,  81 

—  calculation,  copper,  73 

—  in  pig  iron,  193 

—  roasting  of  lead  ores,  125 
Sluice  amalgamation,  silver,  38 
Smelting,  electric,  for  iron,  207 
— ,  — ,  of  nickel  matte,  174 

—  for  copper  matte,  71,  87 

—  for  zinc  in  the  Rhenish  provinces, 
239 

—  of  bottoms,  nickel,  165 

—  of  chrome  iron  ore,  218 

—  of  low-grade  silver  ore,  29 

—  of  oxidized  tin  ores,  135 

—  of  poor  tin  slags,  141 

—  of  rich  tin  slags,  140 

—  of  tops,  nickel,  165 

— ,  reaction,  for  lead,  119 
— ,  reducing  of  iron  ores,  180 
— ,  reduction,  for  lead,  123 
—  roast-reaction,  for  nickel,  170 
Solution  and  precipitation  of  gold,  21 
— ,  chemical,  of  gold,  2 


Solution,  chemical,  of  platinum,  25, 

26 

Solution,  chlorinating,  of  gold,  21 
— ,  electrolytic,  of  tin,  141 

—  of  gold  in  metals,  1 

— •  of  silver  in  chemical  solvents,  47 
,  in  metals,  28 

—  of  platinum,  25 
Sorbite,  193 

Spence  roasting  furnace,  69 

Spiegeleisen,  196,  251 

Stadtberge  works,  leaching  of  copper 

ores  at,  91 

Stall  roasting  of  copper  ores,  63 
Stalmann  converter,  96,  98,  99 
Stamp-mill  amalgamation,  38 
Starring  antimony,  155 
Steel,  197 

— ,  alloying  in  the  bath,  215 
— ,  case  hardening,  216 

—  cementation,  216 

— ,  electric  smelting,  207 
— ,  Girod  furnace,  209,  210 
— ,  Heroult  furnace,  208 

—  induction  furnaces,  Colby,  211 

,  de  Ferranti,  211 

,  Kjellin,  211 

,  Rochling-Rodenhauser,  212, 

213 

— ,  production  of,  201 
— ,  welding,  215 
Stetefeldt  furnace,  47,  65 
Sticht,  copper  converters  used  by,  96, 

98 

— ,  shaft  furnace,  78,  79 
Stoves  for  heating  blast,  182 
Sublimation  of  antimony,  150 
Sulphur,  melting  bismuth  with,  116 
— ,  removal  from  tin  ores,  134 
Sulphuric  acid  parting  of  silver  and 

gold,  51 

Sulphurizing  fusion,  antimony  refin- 
ing, 155 
,  Rossler's,  51 


Talbot  process,  iron,  203 


INDEX 


271 


Tarnowitz  process,  lead,  121 
Temper  carbon,  193 
Thiel  process,  iron,  203 
Thomas-Gilchrist  process,  iron,  202 
Tin,  133 

— ,  blast  furnace  process,  136 
— ,  chemical  concentration,  133 
— ,  Chinese  furnace  for  smelting  ore, 
137 

—  concentration  of  ores,  133 

—  crystallization,  144 

— ,  electric  furnace,  Borchers  and 
Mattonet,  138 

— ,  electrolytic  solution  and  depo- 
sition, 141 

—  liquation,  143 

— ,  mechanical  concentration,  133 

—  ores,  133 

— ,  oxidizing  fusion,  144 

—  properties,  145 

— ,  reducing  fusion,  135 
— ,  removal  of  sulphur  and  arsenic,  134 
— ,  —  of  tungstates,  134 
— ,  reverberatory  furnace  for  smelt- 
ing, 136,  139 

— ,  smelting  of  poor  slags,  141 
— ,  —  of  rich  slags,  140 
— ,  —  operations  for  oxidized  ores,  135 
Tin-plate  waste,  working  up,  141 
Tops,  smelting  of,  for  nickel,  165 
Troostite,  193 

Tungstates,  removal  from  tin  ores,  134 
Tungsten,  225 
— ,  lixiviation,  226 

—  ores,  225 

— ,  properties,  227 

— ,  reducing  fusion,  225,  227 

— ,  separation   of   iron,    manganese, 

and  calcium  from,  226 
Tungstic  acid,  precipitation  of,  226 
— ,  reduction  of,  227 

W 

Walker  casting  machine,  copper,  103 
Warlimont's  process,  nickel,  168 
Washing  with  acids,  mercury,  59 
Washoe  process,  silver,  46 


Water-jacket  furnace,  lead,  129 
Weidtmann,     constitution     of    leap 

matte,  124 
Welding  iron,  215 
Wethey  furnaces,  copper,  67,  69 
Wet-grinding  amalgamation,  silver,  38 
Wet-mill  amalgamation,  40 
Wet  tube  mill,  gold,  13 
White  iron,  190 

White-Howell  furnace,  copper,  70 
Wolframite,  225 

Wrought  iron,  production  of,  199 
Wiist,    constitution    of    iron-carbon 

alloys,  188,  192 
— ,  malleablizing,  198 

U 
Upper  Harz  lead  furnace,  128 


Vacuum  filter  for  gold  slimes,  19 
Vat  with  filter  bottom  and  Butters' 

distributer,  10 

Vitriolization  process,  Freiberg,  51 
,  Harz,  53 


Ziervogel  process,  silver,  47 

Zinc,  231 

— ,  Belgian  process,  236 

—  deposition,  electrolytic,  246 
— ,  desilverization,  31 

—  distillation,  245 

— ,  electrolytic  deposition,  246 

—  furnaces,  Broken  Hill,  247 
— ,  Faber  du  Faur,  247 
— ,  Schrubko,  237,  238 

— ,  liquating,  245 

— ,  lixiviation,  Hopfner,  246 

—  muffle  system,  235 

—  ores,  231 

— ,  precipitation  of  gold  by,  16 

—  properties,  248 

—  retort  press,  243 

— ,  Rhenish  process,  239 

— ,  roast-reduction  work,  232 

— ,  Silesian  muffles,  235 


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14 


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16 


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Douglas's  Untechnical  Addresses  on  Technical  Subjects 12mo,  1  00 

Eakle's  Mineral  Tables 8vo,  1  25 

Eckel's  Stone  and  Clay  Products  Used  in  Engineering.      (In  Preparation.) 

Goesel's  Minerals  and  Metals:  A  Reference  Book 16mo,  mor.  3  00 

*  Groth's  The  Optical  Properties  -of  Crystals.     (Jackson.) 8vo,  3  50 

Groth's  Introduction  to  Chemical  Crystallography  (Marshall) 12mo,  1  25 

*  Hayes's  Handbook  for  Field  Geologists ,  16mo,  mor.  1  50 

Iddings's  Igneous  Rocks 8vo,  5  00 

Rock  Minerals 8vo,  5  00 

Johannsen's  Determination  of  Rock-forming  Minerals  in  Thin  Sections.  8vo, 

With_Thumb  Index  5  00 

*  Martin's  Laboratory    Guide    to    Qualitative    Analysis    with    the    Blow- 

pipe  c 12mo,  60 

Merrill's  Non-metallic  Minerals:  Their  Occurrence  and  Uses 8vo,  4  00 

Stones  for  Building  and  Decoration 8vo,  5  00 

*  Penneld's  Notes  on  Determinative  Mineralogy  and  Record  of  Mineral  Tests. 

8vo,  paper,  50 
Tables  of  Minerals,   Including  the  Use  of  Minerals  and  Statistics  of 

Domestic  Production 8vo,  1  00 

*  Pirsson's  Rocks  and  Rock  Minerals 12mo,  2  50 

*  Richards's  Synopsis  of  Mineral  Characters 12mo,  mor.  1  25 

*  Ries's  Clays:  Their  Occurrence,  Properties  and  Uses 8vo,  5  00 

*  Ries  and  Leighton's  History  of  the  Clay-working  Industry  of  the  United 

States 8vo,  2  50 

*  Rowe's  Practical  Mineralogy  Simplified 12mo,  1  25 

*  Tillman's  Text-book  of  Important  Minerals  and  Rocks 8vo,  2  00 

Washington's  Manual  of  the  Chemical  Analysis  of  Rocks 8vo,  2  00 

MINING. 

*  Beard's  Mine  Gases  and  Explosions Large  12mo,  3  00 

*  Crane's  Gold  and  Silver 8vo,  5  00 

*  Index  of  Mining  Engineering  Literature 8vo,  4  00 

*  8vo,  mor.  5  00 

*  Ore  Mining  Methods 8vo,  3  00 

Dana  and  Saunders's  Rock  Drilling.      (In  Press.) 

Douglas's  Untechnical  Addresses  on  Technical  Subjects 12mo,  1  00 

Eissler's  Modern  High  Explosives 8vo,  4  00 

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